Gold Project 2011

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Evaluation and

exploitation of gold

ore deposits at el Sukari area in eastern

desert of Egypt

By

B.Sc. Mining Eng. Project

2010/2011

A Graduate Project

Submitted in Partial Fulfillment of the Requirements for the

B.Sc. In Mining Engineering.

Department of Mining Engineering

Faculty of Petroleum and Mining Engineering

Suez Canal University

2011

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Submitted By:

Marketing

Ragab eid said

Sameh karam Hassan

Ore evaluation

Mohamed Mahmoud Mahmoud Afify

Mahmoud Mohamed Ahmed Eliwa

Opening up

Ahmed Ali Mohamed Ali

Hossam Eldin Hassan Ali

Drilling & Blasting

Ahmed Gamal Ahmed Hekal

Mohamed Ibrahim Mostafa

Design of transport system

Ibrahim Gharieb Ibrahim Dawood

Ragab Abd Elaziz Ibrahim Afify

Processing of R.O.M

Amir Mohamed Ahmed

Ahmed Hassanin Ahmed

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Supervised By:

Prof. Dr. Ali Hemeda Gomaa.

Prof. Dr. Mohamed Abd El Tawab El Gendy.

Prof. Dr. Saeed Abd Allah Mohamed.

Prof. Dr. Montaser Sabbah El Dein El Salmawy.

Prof.Dr. Mostafa Abas Hamam

Prof.Dr. Mohamed Hussien Allam.

Prof.Dr.Ahmed sedik

Dr. Abd El Azeem Mahmoud Abd El Aal.

Dr. Khaled Fayez

Prof.Dr. Salah Sameeh

Eng. Abd el menaem Selem

Eng. Amr Fathy

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ACKNOWLEDGEMENT

We have the great pleasure to express our deep appreciation and

thanks to all staff members of mining and minerals engineering

department, Suez Canal University Especially for Prof. Dr./

Mohamed Abd El Tawab El Gendy, Prof. Dr. \Saeed Abd Allah

Mohamed and Dr. Abd El Azeem Mahmoud Abd El Aal for their

advice and helpful discussion.

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Table of Contents

Geography .......................................................................................... - 7 -

Geology .............................................................................................. - 7 -

Marketing ......................................................................................... - 16 -

Ore Reserve Estimation ................................................................... - 44 -

Opening up ....................................................................................... - 99 -

Drilling and balsting ........................................................................ - 99 -

Loading and transportation ............................................................ - 145 -

Ore dressing ................................................................................... - 157 -

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Geography

The historic Sukari gold mine is at about 24°57'N 34°43'E, about 20 km from the

Red Sea, and 10 km south of a sealed

road that runs from Edfu on the Nile

to Marsa Alam on the Red Sea.

There is a recently opened

international airport about 70 km

north of Marsa Alam. Five-star

tourist resorts are now springing up

along the Red Sea coast for many

kilometres, almost overnight.

Geology

Theories of gold genesis:

Gold is concentrated by various

natural geologic processes to form commercial deposits of two principal types: hard rock

lode (primary) deposits and placer (secondary) deposits.

There three theories of gold genesis :-

1. Lode deposits are the gold deposits which remain locked within their original

solid rock formations. These are the targets for the "hard rock" prospector

seeking gold at the site of its deposition which was formed from mineralizing

solutions within the earth.

2. Another model which applies to the origins of some gold deposits, suggests

that gold-bearing solutions may be expelled from magma as it cools,

precipitating ore materials as they move into cooler surrounding rocks.

3. A third theory of gold genesis is applied mainly to gold-bearing veins in

metamorphic rocks that occur in mountain belts at continental margins.

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Gold-bearing minerals

Calaverite AuTe2, 44.5% gold; sylvanite (―graphic tellurium‖), AuTe3; native gold

alloyed with silver, etc

Gold most commonly occurs in quartz veins both as native, and as scales and wires

mechanically mixed in pyrite. It may be set free by oxidation and removal of pyrite.

It also accompanies mispickel, chalcopyrite, and rarely galena. On oxidizing and losing

tellurium, they yield extremely fine particles, not readily panned and resisting

amalgamation: called ―rusty‖ gold.

Predominant gold

A. Fissure veins containing native gold, alone, or mechanically mixed in pyrite and

much rarer base-metal sulphides, in quartz gangue. Gray, greasy-looking quartz seems

to accompany best values. Veins appear most frequently in schists, slates, or other

metamorphic rocks, and in association with intrusive rocks, of which granite is

commonest.

B. Impregnations and replacements of open-textured rocks with gold-bearing pyrite.

The ‖banket‖ of gold-bearing conglomerates of Transvaal (in South Africa) is the

best example.

C. Saddle reefs or arch-like deposits of gold-bearing quartz at crests of anticlines.

Saddle-reefs may succeed one another in depth. Slates or slaty schists are common

wall-rocks.

D. Veins carrying gold tellurides. We are associated with an eroded Eocene volcano,

often favoring neighborhood of minor dikes of phonolite and basaltic rocks, with which

volcanic activity closed. Purple fluorite is a characteristics associate.

E. Lateral impregnations and replacements of calcareous shales, with tellurides

along supply fissures, called verticals.

F. Contact zones, on the border of intrusive igneous rock and limestone, containing

gold-bearing mispickel in lime silicates. The usual contact zone of this type carries

copper sulphides with a little gold.

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G. Placer deposits of gold-bearing gravels, which may be: residual, from weathering

of rocks in situ; alluvial fans; sea-beaches with active surf; sea-beaches now elevated

and inland. Gold in streams favors places where current has been checked, as the inside

of bends; junctions of tributaries; heads of quiet reaches.

Sukari geology

The vein-type deposit is hosted in Late Neoproterozoic granite that intruded

island-arc and ophiolite rock assemblages. The vein-forming process is related to overall

late Pan-African shear and extension tectonics. At Sukari, bulk NE– SW strike-slip

deformation was accommodated by a local flower structure and extensional faults with

veins that formed initially at conditions of about 300 ºC and 1.5–2 kbar. Gold is

associated with sulfides in quartz veins and in alteration zones. Pyrite and arsenopyrite

dominate the sulfide ore beside minor sphalerite, chalcopyrite and galena. Gold occurs in

three distinct positions: (1) anhedral grains (GI) at the contact between As-rich zones

within the arsenian pyrite; (2) randomly distributed anhedral grains (GII) and along

cracks in arsenian pyrite and arsenopyrite, and (3) large gold grains (GIII) interstitial to

fine-grained pyrite and arsenopyrite. Fluid inclusion studies yield minimum

veinformation temperatures and pressures between 96 and 188ºC, 210 and 1,890 bar,

respectively, which is in the range of epi- to mesothermal hydrothermal ore deposits.

The structural evolution of the area suggests a longterm, cyclic process of repeated

veining and leaching followed by sealing, initiated by the intrusion of granodiorite.

General geology of the central Eastern Desert

The Neoproterozoic crust in NE Africa was consolidated by accretion of

intra-oceanic island arcs, continental micro-plates and oceanic plateaus. As a result of

convergent tectonics, a nappe assembly with two major tectonostratigraphic units was

established within the central Eastern Desert of Egypt: The structural basement, referred

to as ‗‗infrastructure’’ consists of orthogneisses, psammitic schists and amphibolites that

suffered amphibolite grade, polyphase metamorphic conditions. Structural cover units

summarized here as Pan-African Nappes and referred to as suprastructure‘‘ include

ophiolites, mélange lange-like sediments of accretionary-wedge type and calc-alkaline

volcanics similar to igneous rocks at modern arcs.

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Metamorphism in these units is of greenschist metamorphic grade. Exhumation of

previously buried high-grade structural basement units was achieved by combined

sinistral strike-slip faults and related north- and south-dipping, NE-trending normal

faults. This orogen-scale fault system is known as the Najd Fault System that

accommodated bulk NW–SE extension in the Arabian Nubian Shield. Overall NW–SE

extension exposed ‗‗gneissic domes‘‘ namely the Meatiq-Sibai- and Hafafit Domes (Fig.

1).

"Fig. 1 General geology and location of gold deposits of the central Eastern Desert. Note concentration of

metal occurrences close to major faults. Location of the Najd Fault System (NFS) is marked in the inset"

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Some of these domes have been interpreted as core complexes. Simultaneously with

exhumation of basement rocks, intramontane molasse basins with sediments delivered

from basement domes and PanAfrican Nappes were deposited. In addition, various

syn-tectonic and post-tectonic granitoids have been emplaced during late Pan-African

extension that softened the crust by enhanced advective heat supply.

Geology of the Sukari gold mine area

The mine occurs within a Late Neoproterozoic that intruded older

volcanosedimentary successions and an ophiolitic assemblage, both known as Wadi

Ghadir mélange. The volcanosedimentary succession is composed of andesites, dacites,

rhyodacites, tuffs and pyroclastics. Magmatic rocks are of calc-alkaline affinity and

were formed in an island-arc setting. The dismembered ophiolitic succession is

represented by a serpentinite at the base, followed upwards by a metagabbro-diorite

complex and sheeted dykes. Metagabbro-diorite rocks and serpentinites form lenticular

bodies (1–3 km2) as well as small bodies occur conformably scattered in the

volcanosedimentary arc assemblage. All rocks are weakly metamorphosed (lower

greenschist metamorphic facies), intensely sheared and transformed into various schists

along shear zones. Mineralized quartz veins and talc-carbonate veinlets are common.

The Sukari granitoid is elongated in a NNE direction and bounded from west and east by

two steep shear zones (Fig. 2B, C), covering an area of ca. 10 km. The fresh rock is

leucocratic, coarse-grained and pink in color. It has a heterogeneous mineralogical

composition and ranges from monzogranite to granodiorite with dominant quartz,

plagioclase and potash feldspars and less abundant biotite.

In the vicinity of shear zones the granite is foliated, elsewhere, however, it has sharp

intrusive contacts against the older rocks. Along those shear zones serpentinite and

andesite is altered to listvenite rock that attains up to 70 m in thickness and extends for

several kilometers. At the intersection of the two shear zones, where the gold

mineralization is concentrated, the Sukari granite is almost completely altered and

transected by a large amount of quartz veins (Fig. 2B).

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Fig. 2 A Simplified geology and structural frame of the

Sukari mine area. The system of NW-trending sinistral

strike-slip faults (lateral ramps and tear faults) and thrusts

(frontal ramps) forming the arc-shaped structure

developed during NW-directed tectonic thrust (black

arrow). The Sukari granite body intruded the system of

NE-trending sinistral shear zones with locally developed

flower structures. Inset: Flower structure block model of

the Sukari area looking northeast. Steep thrust faults,

strike-slip faults and orientation of major quartz veins

(qu) are indicated. A antithetic shear, T tension gash. B

Map of the Sukari mine area including shear zones and

quartz veins. C NW–SE section (1–2) across the Sukari

granite close to the mine. Shear zones are indicated

(modified from Khalaf and Oweiss 1993).

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Gold ore zone

Gold occurs in two textural positions and three generations in quartz veins or

veinlets: (1) as inclusions in pyrite and arsenopyrite (GI and GII) and (2) as interstitial

grains between pyrite and other

sulfides (GIII). Gold inclusions

(2–20 lm) in pyrite are either

located at the surfaces of As-rich

zones (GI) as revealed by BSE

images (Fig. 3A) or randomly

distributed. Gold inclusions in

arsenopyrite are randomly

distributed or located along

deformational cracks (GII, Fig.

3A). Interstitial gold grains

(GIII, Fig. 7B) are usually

associated with deformed pyrite

and arsenopyrite in the deformed and sheared smoky quartz (type Q2). In this textural

position, gold grains range from 2 to 80 lm and sometimes host small arsenopyrite and

pyrite crystals. Electron microprobe analysis (Table 1) revealed that gold is always

electrum (12–14 wt% silver).

No systematic compositional

difference between inclusion and

interstitial gold were detected.

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Structural controls to gold mineralization

The Sukari gold deposit is a large, sheeted vein-type and brittle-ductile shear zone

hosted gold deposit, developed in a late- to post-orogenic granitoid intrusive complex

intruded into the Neoproterozoic Hijaz Magmatic Arc of the Arabian-Nubian Shield.

Deformation at Sukari is manifest as a fold-thrust-nappe in the foreland to a large

metamorphic core complex - the Hafafit Culmination, uplifted at ca.680Ma and forming

part of the major Najd Fault System. Sukari lies on one of these arcuate, NE trending

thrusts. Gold mineralisation at Sukari (ca.530Ma) postdates the final stages of uplift, so

may not be related to Najd faulting event. Post-accretion shortening direction which

prevailed during the evolution of the Najd Fault System varied between NW-SE and

ESE-WNW.

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Marketing

History of gold

First smelting of gold

Egyptian goldsmiths carry out the first melting or fusing of ores in order

to separate the metals inside. They use blowpipes made from fire-resistant

clay to heat the smelting furnace.

2600 BC

Early gold jewellery

Goldsmiths of ancient Mesopotamia (modern-day Iraq) craft one of the

earliest pieces of gold jewellery, a burial headdress of lapis and carnelian

beads with willow leaf-shaped gold pendants.

1200-1500 BC

Advances in jewellery making

Artisans develop the lost-wax jewellery casting technique. The process

allows for improved hardness and colour variation which in turn broadens

the market for gold products.

1223 BC

Creation of Tutankhamen's funeral mask

Instantly recognised the world over, the funeral mask of Tutankhamun

is a triumph of gold craftsmanship from the ancient world.

600 BC

First gold dentistry practiced

The first use of gold in dentistry as the Etruscans begin securing

substitute teeth with gold wire. Bio-compatibility, malleability and

corrosion resistance still make gold valuable in dental applications.

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564 BC

First international gold currency created

King Croesus develops improved gold refining techniques, permitting

him to mint the world's first standardised gold currency. Their uniform gold

content allows 'Croesids' to become universally recognized and traded with

confidence.

300

First gold nanoparticles

The Romans use gold to colour the Lycurgus Cup. Melting gold powder

into glass diffuses gold nanoparticles throughout which then refract light,

giving the glass a luminous red glow.

1300

Hallmarking practice established

The world's first hallmarking system, scrutinising and guaranteeing the

quality of precious metal, is established at Goldsmith's Hall in London -

where London's Assay Office is still located today.

1370

The Great Bullion Famine begins

During the years 1370-1420, various major mines around Europe

become completely exhausted. Mining and production of gold declines

sharply throughout the region in a period known as 'The Great Bullion

Famine'.

1422

Venice's record year

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The Venice Mint strikes a record 1.2 million gold ducats using 4.26

metric tonnes of gold from Africa and Central Asia. These small coins

prove popular as they are easy to mint and carry plenty of value.

1511

Ferdinand unleashes invasion force

King Ferdinand of Spain proclaims "Get gold, humanely if you can, but

at all hazards, get gold!", launching unprecedented expeditions to the

Americas. Within years, the Inca and Aztec civilizations would be virtually

destroyed by Spanish conquerors.

1717

UK gold standard commences

Britain moves onto a de facto pure gold standard, as the government

links the currency to gold at a fixed rate (establishing a mint price of 77

shillings, ten and a half pennies per ounce of gold).

1803

First gold electroplating practiced

The first recorded experiment in electroplating is carried out by

Professor Luigi Brugnatelli at the University of Pavia. Gold electroplating

ensures improved conductivity, now essential to many 21st century

technologies.

1848

California Gold Rush begins

John Marshall discovers gold flakes while building a sawmill near

Sacramento, California. The greatest gold rush of all time follows as

40,000 diggers flock to California from around the World.

1885

South African Gold Rush begins

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While digging up stones to build a house, Australian miner George

Harrison finds gold ore on Langlaagte farm near Johannesburg. Miners

flock to the region. South Africa will go on to become the source of 40% of

the world's gold.

1885

First Faberge Easter egg crafted

Carl Faberge makes his first gold Imperial Easter Egg for Tsar

Alexander III. Named The Hen Egg, it was commissioned as a gift from the

Tsar to his wife, the Empress Maria Fedorovna, beginning a tradition that

lasts until 1917.

1870-1900

Adoption of gold standard

All major countries other than China switch to the gold standard, linking

their currencies to gold. The practice of bimetallism is abandoned.

1925

Gold standard returns

The UK returns to the gold standard at pre-war parity of $4.86=£1 with

sterling convertible to gold at 77sh 10.5d per standard ounce. This follows

the country's departure from the gold standard six years previously at the

outbreak of World War I.

1933

Roosevelt suspends gold

President Roosevelt suspends US dollar convertibility to gold (gold at

US$20.67/oz). The export of all transactions in, and the holding of gold by

private individuals, is forbidden. Presidential proclamation makes the

dollar convertible again in January 1934 at a new price of $35 per troy

ounce.

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1939

World War II closes gold market

The London gold market is closed on the outbreak of war, as at the

beginning of World War II. The world will later return to a fixed system of

exchange rates, this time with currencies fixed to the dollar and the dollar

convertible into gold.

1944

Bretton Woods conference

The Bretton Woods conference sets the basis of the post-war monetary

system. The US dollar is set to maintain a $35=1 oz gold conversion rate.

Other currencies are fixed in terms of US dollar, thus forming a Gold

Exchange Standard.

1961

First gold bonded microchips

Gold bonding wire is used in microchips engineered at Bell Labs in the

USA. Nowadays literally billions of chips are bonded this way every year,

controlling all manner of indispensible electrical devices.

1961

First gold in space

The first manned space flight uses gold to protect sensitive instruments

from radiation. In 1980, 41kgs of gold is included in space shuttle

construction through brazing alloys, fuel cell fabrication and electrical

contacts.

1967

First South African Krugerrand

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The Krugerrand is introduced in 1967, as a vehicle for private

ownership of gold. This iconic coin is actually intended for circulation as

currency.

1971

Gold window closed

The Bretton Woods system of fixed exchange rates comes to an end as

President Nixon "closes the gold window", suspending US dollar

convertibility to gold. The world enters its present day system of floating

exchange rates.

1985

First gold-based arthritis treatment

Pharmaceutical giant, SmithKline & French, develops Auranofin, a

gold-based drug for the treatment of rheumatoid arthritis. The drug receives

regulatory approval and goes on sale for the first time.

1999

First Central Bank Gold Agreement

The First Central Bank Gold Agreement (CBGA) is agreed. 15

European central banks declare that gold will remain an important element

of their reserves and collectively cap gold sales at 400 tonnes per year over

next five years.

2001

First gold used in heart surgery

Boston Scientific markets the first gold-plated stent (Niroyal) used in

heart surgery. Inserted inside large arteries and veins, such stents act like

scaffolding, propping open the blood vessels to allow adequate flow.

2003

K-gold launched in China

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The World Gold Council creates an entirely new market segment with

the launch of K-gold, the first 18k jewellery in China. The jewellery, in

predominantly white and yellow gold, takes its inspiration from Italian

design.

2004

Launch of SPDR® Gold Shares

The market is transformed by an innovative, secure and easy way to

access the gold market. Six years later SPDR® exceeds $55bn in assets

under management.

2009

Central banks return to buying

In the second quarter of the year, central banks collectively become net

purchasers of gold for the first time in two decades. This reflects a

combination of slowing sales from European banks and growing purchases

by emerging market countries.

2010

Gold price sustains record highs

Fiat currencies are undermined by inflation fears and successive

financial crises. The London pm fix achieves 35 separate successive highs

in the year to date.

2011

Gold in catalytic converters

Gold used in catalytic convertors by a leading European diesel car

manufacturer. The first use of gold in automotive emissions control

The Importance Of Gold In Egypt

One of the oldest elements of earth is the gold, which has been valued

since centuries. Especially the tribe, which found out gold, got many

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advantages out of the things gold brought with itself. This element might

not be as attractive as it is valued, but there are certain characteristics of

gold, which make it different from others.

Not only it is a harmless and comfortable element, gold tends to conduct

electricity quite well. With great convenience and comfort, it can be

transformed in to several shapes, and sizes. The history of the discovery of

gold is embedded in old books, and gives us an overview about it.

There is strong association between gold and Egypt, simply because this

country was one of the few civilizations to discover it. They benefited a

great deal from this metal, and it quickly began the resource of the country.

Not just in the olden days, but in the present day, gold holds the top most

value in the business hub.

It was a medium of exchange with aspect to price for purchasing and

selling commodities. Egypt was one of the countries that also with much

determination followed this concept. However, in Egypt, the role of gold

excelled beyond the boundaries of being used as money.

Some people are known to have worshipped the Sun. The people also

constitute the Egyptian civilization. Sun was often seen as a source of life.

For several civilisations, Gold was closely associated to the sun because of

its yellow and gleaming attributes. The Egyptians perceived gold as the

skin of Gods, especially the RA.

Besides the king, no one else was allowed to wear gold in those times.

Gradually and slowly, priests and royal members were also given the

privilege of wearing gold. The tomb of the king is known as the ―house of

gold‖, and the chamber is made out of gold.

Another amazing feature of gold was that it never rusts, which was

associated with the characteristics of god. The top of pyramids were

usually made out of a mixture of gold, and other metals. Due to its holy and

sacred value, it was often used to engrave the coffins. The mask of

Tutankhamun is an example of gold being used for funerary art. However,

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gold was not easy to dig out, and mining was quite a difficult task even

though there was rich gold present.

Due to the resource of gold present in the region, rivalries began

between Egypt and its bordering countries while mining for gold. The army

forces strictly monitored the convicts who were given the task of gold

mining. Nubia and the Eastern desert were abundant with gold; hence, all

the mining took place there. Looking at the current popularity of jewelers

in Egypt proves that the value and respect for gold is still alive.

Characteristics

• Streak : yellow

• Hardness : 2.5 -3

• Sp.gr :15.5 – 19.3

• Color : gold – yellow to brass yellow

• Luster : metallic

• Cleavage : none

• Fracture: soft

• soluble: in aquaregia and mercury

• crystals: cubic

• similar to : pyrite , chalcopyrite , biotite , markasite

• accompanied by: pyrite , chalcopyrite , sphalerite , magnetite ,

quartiz , tourmaline

Some Uses of gold

Jewelry:

Alloys with lower caratage, typically 22k, 18k, 14k or 10k, contain

higher percentages of copper, or other base metals or silver or palladium in

the alloy.

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Medicine:

Gold alloys are used in restorative dentistry, especially in tooth

restorations, such as crowns and permanent bridges.

Industry:

Gold solder is used for joining the components of gold jewelry by

high-temperature hard soldering or brazing.

Electronics:

The concentration of free electrons in gold metal is 5.90×1022 cm−3.

Gold is highly conductive to electricity, and has been used for electrical

wiring in some high-energy applications. gold has the advantage of

corrosion resistance

Commercial chemistry:

Gold is attacked by and dissolves in alkaline solutions of potassium or

sodium cyanide, to form the salt gold cyanide—a technique that has been

used in extracting metallic gold from ores in the cyanide process. Gold

cyanide is the electrolyte used in commercial electroplating of gold onto

base metals and electroforming.

Investing in Gold Mines:

Many holders of gold store it in form of bullion coins or bars as a

hedge against inflation or other economic disruptions. However, some

economists do not believe gold serves as a hedge against inflation or

currency depreciation

An alternative to buying physical gold bullion is to invest in the shares

of companies that are involved in the exploration and mining of

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gold. Analysts consider this to be a more risky investment, as the chances

of making a large return on your investment will depend both on the price

of gold price and the success of that company. However, in theory as the

price of gold bullion goes up, so does the value of gold mining shares.

As this is a highly technical area, if you are considering this kind of

investment we recommend doing plenty of research on the industry and

speaking to an experienced financial advisor. As with all investments,

financial advisors would encourage a diverse portfolio, so it is prudent to

consider a complimentary investment of gold bullion too. Overall, though

an investment in mining shares can be quite risky, it can provide a very

good return on investment.

World gold productive countries

Rank Country/Region Gold production (kilograms)

World 2,310,000

1 South Africa 272,128

2 China 247,200

3 Australia 247,000

http://en.wikipedia.org/wiki/List_of_countries_by_gold_production



http://en.wikipedia.org/wiki/South_Africa

http://en.wikipedia.org/wiki/People's_Republic_of_China

http://en.wikipedia.org/wiki/Australia





































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4 United States 242,000

5 Peru 203,268

6 Russia 159,340

7 Canada 104,198

8 Mali 85,411

9 Uzbekistan 84,000

10 Ghana 66,205

11 Indonesia 58,773

12 Papua New Guinea 58,349

13 Argentina 44,131

14 Chile 42,100

15 Brazil 40,075

16 Tanzania 39,750

17 Philippines 36,098

18 Mexico 35,899

19 Mongolia 22,561

20 Guinea 16,336

http://en.wikipedia.org/wiki/United_States

http://en.wikipedia.org/wiki/Peru

http://en.wikipedia.org/wiki/Russia

http://en.wikipedia.org/wiki/Canada

http://en.wikipedia.org/wiki/Mali

http://en.wikipedia.org/wiki/Uzbekistan

http://en.wikipedia.org/wiki/Ghana

http://en.wikipedia.org/wiki/Indonesia

http://en.wikipedia.org/wiki/Papua_New_Guinea

http://en.wikipedia.org/wiki/Argentina

http://en.wikipedia.org/wiki/Chile

http://en.wikipedia.org/wiki/Brazil

http://en.wikipedia.org/wiki/Tanzania

http://en.wikipedia.org/wiki/Philippines

http://en.wikipedia.org/wiki/Mexico

http://en.wikipedia.org/wiki/Mongolia

http://en.wikipedia.org/wiki/Guinea

















































































































































































































































































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Gold Production By Country (million of ounces)

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50% of all gold ever produced was produced since 1960

80% of all gold ever produced was produced since 1900

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Top Gold Importers

$10.9 billion (33.1% of top ten

gold importers, up 6% from 2004)

India

$4.4 billion (13.5%, up 10.9%) United

States

$3.9 billion (11.8%, up 11.4%) Turkey

$3.5 billion (10.5%, down 2.8%) Italy

$2.3 billion (6.9%, up 60.1%) Canada

$2.1 billion (6.5%, up 18.5%) Australia

$2 billion (6%, up 71.4%) Thailand

$1.4 billion (4.1%, up 9.2%) Malaysia

$1.3 billion (4.1%, up 12.1%) Japan

$1.2 billion (3.5%, up 17.9%). Germany

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Top Gold Exporters

In 2005,

the following nations exported the most gold by value.

United States $ 5.5 billion (23.2% of top ten gold

exporters, up 25.2% from 2004)

Australia $4.4 billion (18.6%, up 7.1%)

Canada $3.7 billion (15.5%, up 28.4%)

Peru $3.1 billion (12.9%, up 30.1%)

Hong Kong $2.8 billion (11.8%, down 55.9%)

Japan $1.4 billion (6%, up 17.8%)

Germany $841.7 million (3.5%, up 22.7%)

Singapore $764.8 million (3.2%, up 14.2%)

Italy $628.7 million (2.6%, up 41%)

Colombia $627.2 million (2.6%, up 9%).

Italy, Peru, America and Germany had the fastest increasing exports for

gold in 2005. However, while higher gold prices in 2008 may lead to

more demand for gold bullion and gold mining stock investments, gold

exports may well decrease as global trade partners wait for gold prices to

normalize.

- 35 - | P a g e

Prices in last 10 years

prices in last 5 years

- 36 - | P a g e

Prices last year

presence of gold in the eastern desert

we can find the gold deposits in many groups :

1-North group :

It includes 30 sites in between 27 45 and 28 00 latitude32 45 and 33 05

longitudes

2-middle group :

It includes 63 sites in between 24 10 and 26 45 latitude33 15 and 35 20

longitudes

3-southern east group :

It includes 7 sites in between 22 15 and 23 20 latitude And 34 25 and 35

50 longitudes

- 37 - | P a g e

4-southern west group :

It includes 19 sites in between 22 00 and 23 00 latitude

Some sites in details

1-Atoud

Location :

55 km south west of marsa alam

Intersection between 24 58 latitude with 34 40 longitudes

Reserve :

Proven : 8595 tons which have 109 kg gold

Possible : 10895 tons which have 78 kg gold

Probable : 13600 tons which have 238 kg

2-Barramiya

Location : 105 km east of edfu city

Intersection between 25 05 latitude and 33 47 longitudes

Reserve :

It contains 16 million tons ore which have 21 tons gold

3-Hangaliya

Location : 80 km south west of marsa alam

Reserve :

- 38 - | P a g e

478490 tons ore which have at least 478 kg gold

4- sukari

Location :

30 km south east of marsa alam

Reserve :

1.2 million tons ore which contain 2447 kg gold

5- umm ulaygah

Location :

80 km south west of ras bnas

Reserve :

Not calculated yet

6- Umm EL Rus

Location :

80 km south of qusair city

Reserve :

Not calculated yet

7-anait

Location :

160 km south of abo ghsoun

- 39 - | P a g e

Reserve :

83000 tons ore which contain 185 kg gold

8-Umm Ud

Location :


Reserve :

15600 tons ore which contains 354 kg gold

9-Hamash

Location :


the production from the beginning of the project until February 2010 is

65 kg and the plan of 2010 refers to the production of 15000 ounces

10 –Umm higab

Location :

40 km north of Hamash , 35 km south east of Barramiya

11-samut

45 km south east of Barramiya , 35 km south west of umm higab

12-Umm samra

60 km north east of Barramiya

Gold content 0.5 – 12 gm\ton

- 41 - | P a g e

13- abu marawat

Location :

North east abu marawat valley

Reserve :

290,000 tons ore which have 1210 kg gold

14-Hamama

Location :

In the road between Qena and Safaga

Reserve :

Not calculated yet

- 41 - | P a g e

. Gold deposits and occurrences in the Eastern Desert of Egypt (compiled

from Kochine and Basyuni, 1968). (1) Umm mongul; (2) Umm Balad; (3)

Wadi Dib; (4) Fatira; (5) Abu Marawat; (6) Wadi Gasus; (7) Semna; (8)

Gebel Semna; (9) Abu Qarahish; (10) Kab Amiri; (11) Sagi; (12) Gidami;

(13) Hamama; (14) Erediya; (15) Abu Had; (16) Atalla; (17) Rebshi; (18)

Umm Esh; (19) Fawakhir; (20) Hammamat; (21) Umm Had; (22) EL Sid;

(23) Umm Selimat; (24) Hammuda; (25) EL Nur; (26) Kareim; (27) Kab EL

- 42 - | P a g e

Abyad; (28) Tarfawi; (29) Sherm ELBahaari; (30) Zeidum; (31) Wadi

Zeidum; (32) Umm Rus; (33) Sigdit; (34) Talat Gadalla; (35) Abu

Muawaad; (36) Daghbag; (37) EL Hisinat; (38) Bokari; (39) Umm Samra;

(40) Abu Dabbab; (41) Abu Qaria; (42) Umm Saltit; (43) Bezah; (44) Umm

Selim; (45) Barramiya; (46) Dungash; (47) Samut; (48) Umm Hugab; (49)

Urf EL Fahid; (50) Atud; (51) Sukkari; (52) Umm Tundeba; (53)

Hanglaliya; (54) Kurdeman; (55) Sabahia; (56) Umm Ud; (57) Allawi; (58)

Lewewi; (59) Dweig; (60) Hamash; (61) Geli; (62) Qulan; (63) Kab EL

Rayan; (64) Sheialik; (65) AbuRahaya; (66) Wadi Khashb; (67) Umm

Eleiga; (68) Betan; (69) Qurga Rayan; (70) Hutit; (71) Kalib; (72) Kurtunos;

(73) EL Hudi; (74) Hariari; (75) Um Shira; (76) Neqib; (77) Haimur; (78)

The Nile Valley (Block E); (79) Umm Garaiart; (80) Marahib; (81) Atshani;

(82) Murra; (83) Filat; (84) Seiga I; (85) Seiga II; (86) Umm Shashoba; (87)

Abu Fass; (88) Umm Tuyur; (89) Betam; (90) Umm Egat; (91) Kurbiai; (92)

Romit.

- 43 - | P a g e

- 44 - | P a g e

Ore Reserve Estimation

Introduction

Ore reserve estimates are assessments of the quantity and tenor of a mineral that

may be profitably and legally extracted from a mineral deposit through mining and/or

mineral beneficiation. (Examination and evaluation of ore deposites. Generally it

means the determination of the extent and value of mineral deposites.)

Practical consideration of Mineral Resources and Mineral Reserves

Mineral Resource confidence classification should take into account practical

considerations such as drilling, sampling and assay integrity, drill hole spacing,

geological control and continuity, grade continuity, estimation method and block size,

potential mining method and reporting period. Ore Reserve confidence classification

should take into account the confidence classification of the Mineral Resource and

should not include Inferred Resources. Cut-off grades, mining and metallurgical

factors or assumptions, cost and revenue factors, market assessment (where

appropriate) and other risk factors such as environmental, social or political should be

considered by the CP in terms of their impact on confidence in the Ore Reserve

estimate.

Mineral Resource

A concentration or occurrence of material of intrinsic economic interest in or on the

Earth‘s crust in such form, quality and quantity that there are reasonable prospects for

eventual economic extraction. The location, quantity, grade, geological characteristics

and continuity of a Mineral Resource are known, estimated or interpreted from specific

geological evidence and knowledge. Mineral Resources are sub-divided, in order of

increasing geological confidence, into Inferred, Indicated and Measured categories.

Measured Mineral Resource

That part of a Mineral Resource for which tonnage, densities, shape, physical

characteristics, grade and mineral content can be estimated with a high level of

confidence. It is based on detailed and reliable exploration, sampling and testing

- 45 - | P a g e

information gathered through appropriate techniques from locations such as outcrops,

trenches, pits, workings and drill holes. The locations are spaced closely enough to

confirm geological and grade continuity.

Ore Reserve

The economically mineable part of a Measured and/or Indicated Mineral

Resource. It includes diluting materials and allowances for losses, which may occur

when the material is mined. Appropriate assessments and studies have been carried out,

and include consideration of and modification by realistically assumed mining,

metallurgical, economic, marketing, legal, environmental, social and governmental

factors. These assessments demonstrate at the time of reporting that extraction could

reasonably be justified. Ore Reserves are sub-divided in order of increasing confidence

into Probable Ore Reserves and Proved Ore Reserves.

Probable Ore Reserve

The economically mineable part of an Indicated, and in some circumstances, a

Measured Mineral Resource. It includes diluting materials and allowances for losses

which may occur when the material is mined. Appropriate assessments and studies

have been carried out, and include consideration of and modification by realistically

assumed mining, metallurgical, economic, marketing, legal, environmental, social and

governmental factors. These assessments demonstrate at the time of reporting that

extraction could reasonably be justified.

Proved Ore Reserve

The economically mineable part of a Measured Mineral Resource. It includes

diluting materials and allowances for losses which may occur when the material is

mined. Appropriate assessments and studies have been carried out, and include

consideration of and modification by realistically assumed mining, metallurgical,

economic, marketing, legal, environmental, social and governmental factors. These

assessments demonstrate at the time of reporting that extraction could reasonably be

justified.

- 46 - | P a g e

Mineral Resources and Mineral Reserves must be reported on a site by site basis.

Where estimates for both Mineral Resources and Mineral Reserves are reported, for

consistency, a single form of reporting should be used in a report. Appropriate forms

of clarifying statements may be:

The Measured and Indicated Mineral Resources are inclusive of those

Mineral Resources modified to produce the Mineral Reserves,‘ or

The Measured and Indicated Mineral Resources are additional to the Mineral

Reserves.‘

Inferred Mineral Resources are, by definition, always additional to Mineral

Reserves.

RESOURCE ESTIMATION METHODOLOGY

A resource estimate is based on prediction of the physical characteristics of a mineral

deposit through collection of data, analysis of the data, and modeling the size, shape,

and grade of the deposit. Important physical characteristics of the ore body that must be

predicted include:

(1) The size, shape, and continuity of ore zones,

(2) The frequency distribution of mineral grade,

(3) The spatial variability of mineral grade.

These physical characteristics of the mineral deposit are never completely known, but

are inferred from sample data.

The sample data consist of one or more of the following

1. Physical samples taken by drilling, trenching, test pitting, and channel sampling.

2. Measurement of the quantity of mineral in the samples through assaying or other

procedures.

3. Direct observations such as geologic mapping and drill core logging.

- 47 - | P a g e

Estimation of the resource requires analysis and synthesis of these data to develop a

resource model.

Methods used to develop the resource model may include

1. Compilation of the geologic and assay data into maps, reports, and computer

databases.

2. Delineation of the physical limits of the deposit based on geologic interpretation of

the mineralization controls at a reasonable range of mining cutoff grades.

3. Compositing of samples into larger units such as mining bench height, seam

thickness, or minable vein width.

4. Modeling of the grade distribution based on histograms and cumulative frequency

plots of grades.

5. Evaluation of the spatial variability of grade using experimental variograms.

6. Selection of a resource estimation method and estimation of quantity and grade of

the mineral resource.

The estimation procedure must be made with at least minimal knowledge of the

proposed mining method since different mining methods may affect the size, shape,

and/or grade of the potentially minable ore reserve. The most important mining

factors for consideration in evaluation of the ore reserve from the resource are:

The range of likely cutoff grades. The degree of selectivity and the size of the

selective mining unit for likely mining methods. Variations in the deposit that affect the

ability to mine and/or process the ore.

These mining factors often determine the degree of detail that is required for the

resource model and thus the degree of difficulty to develop a resource model for

estimating ore reserves. For example, a disseminated gold deposit may be continuous

and regular in shape, if mined by bulk, open pit methods. The same deposit may be

discontinuous and difficult to estimate, however, if mined by more selective

underground methods at a higher cutoff grade. Such large differences in deposit shape

due to variations in cutoff grade and mining method may require different ore reserve

estimation methods or different mining methods.

- 48 - | P a g e

DATA COLLECTION AND GEOLOGIC INTERPRETATION

Data that must be collected and compiled for the resource estimate are as follows:

1. Reliable assays from an adequate number of representative samples.

2. Coordinate locations for the sample data.

3. Consistently recorded geologic data that describe the mineralization controls.

4. Cross sections or plan maps with the geologic interpretation of the mineralization

controls.

5. Tonnage factors or specific gravities for the various ore and waste rock categories.

6. Surface topographic map, especially for deposits to be surface mined.

Although small deposits may be evaluated manually using data on maps and in reports,

the amount of data required for a resource estimate is often large, and data may be more

efficiently evaluated if they are entered into a computer database. Computer programs

can then be used to retrieve the data for printing reports, plotting on digital plotters,

statistical analysis, and resource estimation. Minimum information that should be in-

cluded in a drillhole database are :

1. Drillhole number or other identification.

2. Hole length, collar coordinates, and down-hole surveys.

3. Sample intervals and assay data.

4. Geologic data such as lithology, alteration, oxidation, etc.

5. Geotechnical data such as RQD (rock quality designation).

Types of reserve calculation

The classical, employing two-dimensional maps and hand calculation. The

geostistical, a more soghisticaded approach requiring digital computers to prepare

statistically derived estimates. Different classical procedures are used in calculating

reserve estimates. They differ mainly in the ways in which they combine the sample

data.

- 49 - | P a g e

Reserve estimation from maps methods

1. polygon method

2. The triangle method

3. In the section method

polygon method

In the section method

- 51 - | P a g e

The triangle method

Types of manual methods

1. Vertical sections

2. Horizontal sections

3. Block module

The method used to calculation

Vertical section:

Av. Assay for each bore hole:

Av. Au = ∑ (D x Au) / ∑ D

Av. Assay for every section:

Av. Au = ∑ (W x D x Au)/ ∑ (W x D)

Procedures:

The procedure for the calculation of a reserve estimate – tonnage and average

grade of ore for a mineral deposit is demonstrated in the case study.

- 51 - | P a g e

Sec. 10000 mN

Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU

D1321 10000 10965 3 1.11 3.33 1.08 35 105 116.55

sum

9 1.07 9.63 35 315 337.05

12 12.96 420 453.6


RCD99 10000 371.3 40 2.86 114.4 3.82104167 35 1400 4004

1 12.2 12.2 35 35 427

6 8.09 48.54 35 210 1698.9

1 8.27 8.27 35 35 289.45

sum 48 183.41 1680 6419.35


RCD416 10000 10523 23 6.92 159.16 8.12410256 35 805 5570.6

sum

11 12.53 137.83 35 385 4824.05

5 3.97 19.85 35 175 694.75

39 316.84 1365 11089.4


RCD1173 10000 10675 26 3.59 93.34 4.11560606 35 910 3266.9

sum

1 27.3 27.3 35 35 955.5

1 35.7 35.7 35 35 1249.5

- 52 - | P a g e

33 2.37 78.21 35 1155 2737.35

1 8.75 8.75 35 35 306.25

1 9.32 9.32 35 35 326.2

2 6.27 12.54 35 70 438.9

1 6.47 6.47 35 35 226.45

66 271.63 2310 9507.05

sec 10000N Sum 5775 27469.4

average assay 4.756606061

- 53 - | P a g e

- 54 - | P a g e

Sec. 10050 mN


D1313 10050 10665 45 1.07 48.15 1.07 50 2250 2407.5

sum 45 48.15 2250 2407.5


D1388 10050 10550 3 1.54 4.62 9.96142857 50 150 231

sum

20 7.39 147.8 50 1000 7390

5 25.3 126.5 50 250 6325

28 278.92 1400 13946


RCD118

9 10050 10560 20 2.46 49.2 2.46 50 1000 2460

sum 20 49.2 1000 2460


RCD126

3 10050 10625 2 4.1 8.2 3.88 50 100 410

sum

65 3.96 257.4 50 3250 12870

3 2 6 50 150 300

70 271.6 3500 13580

sec 10050N Sum 8150 32393.5


- 55 - | P a g e

- 56 - | P a g e

Sec. 10100 mN


D1328 10100 10660 19 8.03 152.6 4.760714 50 950 7628.5

sum

7 20.58 144.1 50 350 7203

72 2.36 169.9 50 3600 8496

98 466.6 4900 23328


RCD1166 10100 10611 23 6.23 143.3 2.986923 50 1150 7164.5

68 1.89 128.5 50 3400 6426

91 271.8 4550 13591


RCD1174 10100 10665 104 1.59 165.4 1.713396 50 5200 8268

sum

2 8.13 16.26 50 100 813

106 181.6 5300 9081


RCD1225 10100 10662 83 1.95 161.9 2.612667 50 4150 8092.5

sum

1 5.62 5.62 50 50 281

3 15.83 47.49 50 150 2374.5

2 5.75 11.5 50 100 575

1 8.68 8.68 50 50 434

90 235.1 4500 11757

- 57 - | P a g e

sec 10100 Sum 19250 57756


- 58 - | P a g e

- 59 - | P a g e

Sec. 10150 mN


D1289 10150 10725 4 1.62 6.48 2.045 50 200 324

Sum

13 2.17 28.21 50 650 1410.5

1 8 8 50 50 400

8 1.31 10.48 50 400 524

26 53.17 1300 2658.5


RCD500 10150 10556 56 2.03 113.68 2.03 50 2800 5684

Sum 56 113.68 2800 5684


RCD1165 10150 10662 45 2.35 105.75 2.35 50 2250 5287.5

Sum 45 105.75 2250 5287.5

sec 10150N Sum 6350 13630


- 61 - | P a g e

- 61 - | P a g e

Sec. 10200 mN

Hole N E D(m) Au

(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU

D1383 1020

0

1068

5 3 2.82 8.46 2.82 50 150 423

Sum 3 8.46 150 423

Hole N E D(m) Au


D1880 1020

0

1069

0 17 29.7 504.9 51.9421053 50 850 25245

sum

2 241 482 50 100 24100

19 986.9 950 49345

Hole N E D(m) Au


RD502 1020

0

1055

7 38 6.86

260.6

8 9.2288 50 1900 13034

sum

9 19.6

1

176.4

9 50 450 8824.5

3 8.09 24.27 50 150 1213.5

50 461.4

4 2500 23072

Hole N E D(m) Au


RCD11

61

1020

0

1064

9 23 2.6 59.8 3.76884615 50 1150 2990

- 62 - | P a g e

sum

3 12.7

3 38.19 50 150 1909.5

26 97.99 1300 4899.5

Hole N E D(m) Au


RCD11

77

1020

0

1070

6 65

18.3

1

1190.

2 31.9614286 50 3250 59507.5

sum

5 209.

4

1047.

2 50 250 52357.5

70 2237.

3 3500 111865

sec 10200N Sum 8400 189605


- 63 - | P a g e

- 64 - | P a g e

Sec. 10250 mN


D1351 10250 10678 34.4 6 206.4 6.3673165 50 1720 10320

Sum

1.6 85.88 137.408 50 80 6870.4

2 15.95 31.9 50 100 1595

11 3.91 43.01 50 550 2150.5

2 14.4 28.8 50 100 1440

1 9.98 9.98 50 50 499

26 1.51 39.26 50 1300 1963

1 6.26 6.26 50 50 313

79 503.018 3950 25150.9


D1386 10250 10595 37 1.04 38.48 8.3125333 50 1850 1924

sum

5 1.08 5.4

50 250 270

7 1.25 8.75 50 350 437.5

3 6.03 18.09 50 150 904.5

22 14.26 313.72 50 1100 15686

1 239 239 50 50 11950

75 623.44 3750 31172


DGT408 10250 10429 26 4.23 109.98 4.8957143 50 1300 5499

sum 2 13.55 27.1 50 100 1355

- 65 - | P a g e

28 137.08 1400 6854


RCD1162 10250 10700 61 2.27 138.47 2.9880882 50 3050 6923.5

sum

2 5.48 10.96 50 100 548

1 13.6 13.6 50 50 680

4 10.04 40.16 50 200 2008

68 203.19 3400 10159.5

sec 10250N sum 12500 73336.4

avg assay 5.866912

- 66 - | P a g e

- 67 - | P a g e

Sec. 10300 mN


D340 10300 10620 19 5.85 111.15 5.85 50 950 5557.5

Sum 19 111.15 950 5557.5


D1304 10300 10831 15 2.88 43.2 3.3413793 50 750 2160

sum

1 25 25

50 50 1250

12 1.74 20.88 50 600 1044

1 7.82 7.82 50 50 391

29 96.9 1450 4845


RC415 10300 10445 10 3.64 36.4 2.96 50 500 1820

sum

17 2.56 43.52 50 850 2176

27 79.92 1350 3996


RCD388 10300 10644 30 3.65 109.5 11.994925 50 1500 5475

sum

3 20.8 62.34

50 150 3117

24 9.7 232.8 50 1200 11640

1 185 185 50 50 9250

9 23.8 214.02 50 450 10701

- 68 - | P a g e

67 803.66 3350 40183


RCD1125 10300 10700 49 1.86 91.14 2.4228 50 2450 4557

sum

1 30 30 50 50 1500

50 121.14 2500 6057


RCD1178 10300 10758 105 4.1 430.5 5.4367033 50 5250 21525

sum

17 12.5 211.65

50 850 10583

2 76.4 152.7 50 100 7635

3 7.49 22.47 50 150 1123.5

1 14 14 50 50 700

51 2.04 104.04 50 2550 5202

1 30.6 30.6 50 50 1530

2 11.8 23.52 50 100 1176

182 989.48 9100 49474

sec 10300N sum 18700 110113


- 69 - | P a g e

- 71 - | P a g e

Sec. 10350 mN

Hole N E D(m) Au

(g/t)

AU *

D

Avg hole

AU.

DIST.BET

holes w*D w*D*AU

D1284 10350 10753 11 2.98 32.78 2.0438554 50 550 1639

sum

1 6.71 6.71

50 50 335.5

1 6.72 6.72 50 50 336

5 1.9 9.5 50 250 475

35 1.58 55.3 50 1750 2765

1 6.55 6.55 50 50 327.5

1 9.09 9.09 50 50 454.5

26 1.13 29.38 50 1300 1469

1 5.03 5.03 50 50 251.5

1 8.58 8.58 50 50 429

83 169.64 4150 8482

Hole N E D(m) Au

(g/t)

AU *

D

Avg hole

AU.

DIST.BET

holes w*D w*D*AU

D1345 10350 10730 94 3.98 374.12 3.3041036 50 4700 18706

sum

3 27.84 83.52

50 150 4176

1 5.18 5.18 50 50 259

4 11.56 46.24 50 200 2312

1 38 38 50 50 1900

1 12.5 12.5 50 50 625

7 10.52 73.64 50 350 3682

54 1.64 88.56 50 2700 4428

- 71 - | P a g e

4 9.62 38.48 50 200 1924

3 2.13 6.39 50 150 319.5

43 0.87 37.41 50 2150 1870.5

15 0.72 10.8 50 750 540

21 0.69 14.49 50 1050 724.5

251 829.33 12550 41466.5

Hole N E D(m) Au

(g/t)

AU *

D

Avg hole

AU.

DIST.BET

holes w*D w*D*AU

D1366 10350 10640 58 2.53 146.74 10.348934 50 2900 7337

sum

2 23.34 46.68

50 100 2334

1 27.4 27.4 50 50 1370

1 6.86 6.86 50 50 343

2 6.72 13.44 50 100 672

10 45.94 459.4 50 500 22970

2 227.7 455.4 50 100 22770

39 1.76 68.64 50 1950 3432

7 5.43 38.01 50 350 1900.5

122 1262.6 6100 63128.5

Hole N E D(m) Au

(g/t)

AU *

D

Avg hole

AU.

DIST.BET

holes w*D w*D*AU

RCD112

4 10350 10700 42 6.24 262.08 8.9569444 50 2100 13104

sum 1 213 213 50 50 10650

- 72 - | P a g e

27 3.66 98.82 50 1350 4941

1 26 26 50 50 1300

1 45 45 50 50 2250

72 644.9 3600 32245

sec 10350 sum 26400 145322


- 73 - | P a g e

- 74 - | P a g e

Section10350


D1143 10400 10710 212 2.37 502.44 3.3339631 50 10600 25122

sum

1 8 8

50 50 400

1 94.2 94.2 50 50 4710

3 39.61 118.83 50 150 5941.5

217 723.47 10850 36173.5


D1307 10400 10845 19 5.31 100.89 8.315 50 950 5044.5

2 9.27 18.54 50 100 927

1 63.5 63.5 50 50 3175

sum 22 182.93 1100 9146.5


RCD523 10400 10696 6 5.09 30.54 5.9487952 50 300 1527

1 15.8 15.8

50 50 790

9 4.07 36.63 50 450 1831.5

0.6 26.3 15.78 50 30 789

sm 16.6 98.75 830 4937.5


- 75 - | P a g e

RCD1186 10400 10749 36 1.91 68.76 2.77625 50 1800 3438

sum

1 5.32 5.32

50 50 266

1 5.87 5.87 50 50 293.5

1 6.3 6.3 50 50 315

1 16.1 16.1 50 50 805

16 3.32 53.12 50 800 2656

56 155.47 2800 7773.5

sec 10400N sum 15580 58031


- 76 - | P a g e

- 77 - | P a g e


D338A 10450 10506 5 14.76 73.8 14.76 50 250 3690

sum 5 73.8 250 3690


D1073 10450 10751 197 2.18 429.46 2.8535149 50 9850 21473

sum

3 12.35 37.05

50 150 1852.5

2 54.95 109.9 50 100 5495

202 576.41 10100 28820.5


RCD508 10450 10636 60 2.4 144 2.4 50 3000 7200

sum 60 144 3000 7200


RCD1187 10450 10750 51 4.45 226.95 6.6868421 50 2550 11347.5

sum

1 36 36

50 50 1800

9 15.51 139.59 50 450 6979.5

14 4.44 62.16 50 700 3108

1 43.5 43.5 50 50 2175

76 508.2 3800 25410

- 78 - | P a g e


RCD1187 10450 10742 4 5.84 23.36 1.846405 50 200 1168

sum

1 20 20

50 50 1000

2 10.52 21.04 50 100 1052

3 1.41 4.23 50 150 211.5

38.4 0.54 20.736 50 1920 1036.8

48.4 89.366 2420 4468.3


RCD1235 10540 10800 8 4.7 37.6 7.0777778 50 400 1880

sum

1 26.1 26.1 50 50 1305

9 63.7 450 3185

sec 10450N sum 20020 72773.8


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D1042 10500 10767 75 2.17 162.8 2.784937 50 3750 8138

sum

1 7.37 7.37

50 50 368.5

3 16.63 49.89 50 150 2495

79 220 3950 11001


D1231 10500 10795 4 2.93 11.72 2.010952 50 200 586

sum

4 3.14 12.56

50 200 628

34 1.77 60.18 50 1700 3009

42 84.46 2100 4223


D1320 10500 10875 2 2.44 4.88 6.288 50 100 244

sum

2 11.24 22.48

50 100 1124

1 4.08 4.08 50 50 204

5 31.44 250 1572


RCD509 10500 10608 34 3.08 104.7 3.08 50 1700 5236

sum 34 104.7 1700 5236


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RCD1213 10500 10500 28 2.56 71.68 4.865902 50 1400 3584

sum

3 10.75 32.25

50 150 1613

26 4.28 111.3 50 1300 5564

1 18.7 18.7 50 50 935

3 20.97 62.91 50 150 3146

61 296.8 3050 14841

sec 10500N sum 11050 36873


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D1013 10550 10700 43 1.61 69.23 3.17228571 50 2150 3461.5

sum

1 6.81 6.81

50 50 340.5

22 3.83 84.26 50 1100 4213

2 11.73 23.46 50 100 1173

2 19.15 38.3 50 100 1915

70 222.06 3500 11103


D1280 10550 10843 24 0.93 22.32 99.4226087 50 1200 1116

sum

46 2.21 101.66

50 2300 5083

2 6.47 12.94 50 100 647

1 5.71 5.71 50 50 285.5

1 11.3 11.3 50 50 565

3 5.74 17.22 50 150 861

35 164.09 5743.2 50 1750 287158

1 5420 5420 50 50 271000

2 49.65 99.3 50 100 4965

115 11434 5750 571680


D1280 10550 10843 24 0.93 22.32 1.6787931 50 1200 1116

sum 31 1.82 56.42 50 1550 2821

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2 6.46 12.92 50 100 646

1 5.71 5.71 50 50 285.5

58 97.37 2900 4868.5


D1379 10550 10770 4 1.66 6.64 1.86862605 50 200 332

sum

3 1.17 3.51

50 150 175.5

2 1.85 3.7 50 100 185

24 1.08 25.92 50 1200 1296

19 1.95 37.05 50 950 1852.5

84 1.7 142.8 50 4200 7140

1 6.45 6.45 50 50 322.5

1 10.9 10.9 50 50 545

1 7.21 7.21 50 50 360.5

1 8.01 8.01 50 50 400.5

13.3 2.23 29.659 50 665 1482.95

1 6.48 6.48 50 50 324

154.3 288.33 7715 14416.5

sec 10550N 19865 602068


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D382 10600 10558 21 2.2 46.2 2.2 50 1050 2310

sum 21 46.2 1050 2310


RC659 10600 10819 3 1.14 3.42 1.3028 50 150 171

sum

21 1.02 21.42

50 1050 1071

1 7.73 7.73 50 50 386.5

25 32.57 1250 1628.5


RCD391 10600 10832 14 3.93 55.02 4.4514 50 700 2751

sum

35 4.21 147.4

50 1750 7367.5

1 20.2 20.2 0 0

50 222.6 2450 10119

sec 10600N sum 4750 14057


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D350 10650 10612 2 4.8 9.6 2.2063636 50 100 480

sum

9 1.63 14.67 50 450 733.5

11 24.27 550 1213.5


D360 10650 10608 168 2.55 428.4 3.2710545 50 8400 21420

sum

18 3.43 61.74

50 900 3087

10 4.42 44.2 50 500 2210

49 3.95 193.55 50 2450 9677.5

5 10.13 50.65 50 250 2532.5

25 4.84 121 50 1250 6050

275 899.54 13750 44977


D1371 10650 10703 34 1.11 37.74 1.2297143 50 1700 1887

sum

1 5.3 5.3 50 50 265

35 43.04 1750 2152

Hole N E D(m) Au % % * D Avg hole(%) Eff.(w) w*D w*D*%

RCD1279 10650 10818 7 1.25 8.75 9.6926667 50 350 437.5

sum

3 4.6 13.8

50 150 690

3 1.1 3.3 50 150 165

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2 59.77 119.54 50 100 5977

15 145.39 750 7269.5


RCD1279 10650 10618 120 1.63 195.6 1.63 50 6000 9780

sum 120 195.6 6000 9780

sec 10650N sum 22800 65392


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RC677 10700 10415 1 24.3 24.3 24.3 50 50 1215

sum 1 24.3 50 1215


RCD402 10700 10858 8 9.09 72.72 10.733333 50 400 3636

sum

4 14.02 56.08 50 200 2804

12 128.8 600 6440


RCD738 10700 10420 16 1.09 17.44 1.09 50 800 872

sum 16 17.44 800 872

sec 10700N sum 1450 8527


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D351 10750 10590 10 2.18 21.8 3.8235 50 500 1090

Sum

15 3.02 45.3

50 750 2265

2 6.7 13.4 50 100 670

36 2.83 101.9 50 1800 5094

3 6.43 19.29 50 150 964.5

24 1.69 40.56 50 1200 2028

9 8.28 74.52 50 450 3726

1 65.6 65.6 50 50 3280

100 382.4 5000 19118


D356 10750 10589 18 1.06 19.08 1.433415 50 900 954

Sum

17 1.53 26.01

50 850 1300.5

6 2.28 13.68 50 300 684

41 58.77 2050 2938.5


D395 10750 10807 1 573 573 573 50 50 28650

Sum 1 573 50 28650


D648 10750 10490 2 1.6 3.2 1.6 50 100 160

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Sum 2 3.2 100 160


D757 10750 10485 5 8.34 41.7 3.429032 50 250 2085

Sum

1 33 33

50 50 1650

15 0.9 13.5 50 750 675

10 1.81 18.1 50 500 905

31 106.3 1550 5315

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D1331 10750 10925 1 3.41 3.41 6.03 50 50 170.5

Sum

1 8.65 8.65 50 50 432.5

2 12.06 100 603


RC678 10750 10464 8 1.16 9.28 3.675 50 400 464

Sum

4 7.45 29.8

50 200 1490

2 14.15 28.3 50 100 1415

6 1.02 6.12 50 300 306

20 73.5 1000 3675


RCD834 10750 10646 31 2 62 3.045741 50 1550 3100

Sum

4 6.92 27.68

50 200 1384

4 5.77 23.08 50 200 1154

1 20.4 20.4 50 50 1020

13 2 26 50 650 1300

1 5.31 5.31 50 50 265.5

54 164.5 2700 8223.5

sec 10750N 12550 68683


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calculation of average assay and T.F from sections.

Sec SEC.area(A)m² SEC.INTERVAL av. depth Volume w= V / TF Au(g/t) w * Au

10000 5320 152 446.1 2373252 6787501 4.756606 32285467

10050 5750 115 575.725 3310419 9467798 3.974663 37631301

10100 2700 54 588.65 1589355 4545555 3.000312 13638083

10150 8450 169 532.233 4497371 12862482 2.146457 27608761

10200 7450 149 534.66 3983217 11392001 22.57196 2.57E+08

10250 13550 271 466.45 6320398 18076337 4.866921 87976107

10300 19300 386 444.15 8572095 24516192 5.888369 1.44E+08

10350 5656 113 633.125 3580955 10241531 5.504621 56375750

10400 7450 149 525.9 3917955 11205351 3.724711 41736697

10450 15000 300 566.244 8493660 24291868 3.635055 88302273

10500 18750 375 506.18 9490875 27143903 3.336878 90575886

10550 7150 143 637.175 4555801 13029592 30.30798 3.95E+08

10600 13700 274 315.45 4321665 12359962 2.959368 36577681

10650 10000 200 521.733 5217330 14921564 2.82807 42199230

10700 22150 443 336 7442400 21285264 2.88069 61316240

10750 23050 461 331.104 7631947 21827369 5.472709 1.19E+08

Mine 2.44E+08 6.740961 1.53E+09

Tonnage Factor (T.F) = 1/density = 1/2.41= 0.4149

Total Tonnage = 243954267.9 tons

average grade = 1532079086 / 243954267.9 = 6.28019 (gm/ton) say 6.3 (gm/ton)

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Opening up

Introduction

Open pit mining refers to a method of extracting rocks and minerals from the earth by

their removal from pit .

Open pit mines is used when deposit of commercially useful mineral or rock are found

near the surface; that is, where the overburden (surface material covering the valuable

deposit) is relatively thin or the material of interest is structurally unsuitable for

tunneling (as would be the case for sand, cinder, and gravel). For minerals that occur

deep below the surface—where the overburden is thick or the mineral occurs as veins

in hard rock—underground mining methods extract the valued material.

Open-pit mines that produce building materials and dimension stone are commonly

referred to as quarries. People are unlikely to make a distinction between an open-pit

mine and other types of open-cast mines such as quarries, borrows, placers, and strip

mines.

Open-pit mines are typically enlarged until either the mineral resource is exhausted, or

an increasing ratio of overburden to ore makes further mining uneconomic. When this

occurs, the exhausted mines are sometimes converted to landfills for disposal of solid

wastes. However, some form of water control is usually required to keep the mine pit

from becoming a lake.

Generally the mining method in regular ores determines by the stripping ratio that

refers to the ratio between volume of over burden and volume of ore, but in disseminate

cute of grade determine the difference between ore and waste,

Cutoff grade = mining cost / (value X recovery)

The mining method used in this mine is surface mining "selective open pit mining

method because of gold deposit".

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Pit geometry design

Final pit limits that refer to the final limits that ore can be extracted and processed with

a profit affected by:

1. Technical factors,

2. Economic factors.

Pit geometry:

1. Pit overall slope angle,

2. Overall pit depth,

3. Pit width and length.

Pit overall slope angle determination

Slope stability analysis forms an integral part of the opencast mining operations

during the life cycle of the project. In Indian mining conditions, slope design

guidelines were not yet formulated for different types of mining practices and there is a

growing need to develop such guidelines for maintaining safety and productivity. Till

date, most of the design methods are purely based on field experience, rules of thumb

followed by sound engineering judgment. During the last four decades, the concepts of

slope stability analysis have emerged within the domain of rock engineering to address

the problems of design and stability of excavated slopes.

The actual slope angles used in

the mine depend upon:

1. The presence of haulage roads,

or ramps, necessary for the

transportation of the blasted

ore from the pit,

2. Possible blast damage,

3. Ore grades,

4. Economic constraints.

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In slope stability analysis, engineering geological investigation of geological data

collection is necessary to determine the regional geology of the project area and to

obtain the reliable data for proposed project area. There are two investigation stages:

1. Initial site investigation,

2. Final site investigation.

In slope failure analysis, determination of mode of failure is a very important and it

can be considered by the value of slope mass rating (SMR).

Plane failure occurs when a geological discontinuity, such as a bedding plane,

strike parallel to the slope face and dip into the excavation (daylight) at an angle greater

than the angle of friction as shown in Fig6(a).

Wedge failure When two discontinuities strike obliquely across the slope face

and their line of intersection daylight in the slope face, the wedge of the rock resting on

these continuities will slide down the line of intersection, provided that the inclination

of this line significantly greater than the angle of friction as shown in Fig6 (b).

Circular failure When the material is weak (soil) or very heavily fractured (waste

rock dump), the failure will be defined by a single discontinuity surface but will tend to

follow a circular failure as shown in Fig6 (c).

Toppling failure can be occurred when the pit slope angle is greater than the angle

of internal friction and the ratio of block width height is less than the friction angle as

shown in Fig6 (d)

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Fig 6: Types of failure (a) Plane failure, (b) Wedge failure, (c) Circular failure and (d) Toppling failure.

To prevent all modes of failure and to

achieve safety work the overall slope

angle takes to be 55º.

The inter ramp slope angle takes to be

70 º.

Pit depth

Shallow ore deposits are mined by surface methods but a depth is reached in the case of

most deposits after which underground methods are applied for the extraction of the

remaining ore.

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From sections:

Section Pit depth

10000mN 221.05

10050mN 225.7

10100mN 318.18

10150mN 323.8

10200mN 333.3

10250mN 360

10300mN 375

10350mN 375

10400mN 350

10450mN 340

10500mN 310

10550mN 324

10600mN 304

10650mN 307

Average pit depth = 295m

Pit width:

Av.Wb = Av. Wd + (ht/tan φ)

Wd: horizontal thickness of the ore body (m),

φ : pit side slope angle along the foot-wall (deg),

ht : pit depth (m).

Wb = 187.319 +2 (295/tan55º)

=450 m

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Pit length:

Lb = Ld + (ht/tan φ)

Ld : total distance between sections (m),

φ : pit side slope angle along the foot-wall (deg),

ht : pit depth (m).

Lb = 650 + (295/tan55º)

=850 m

Parameter value

Over all slope angle 55 º

Pit depth 295 m

Pit width 320 m

Pit length 850 m

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Bench design

A bench may be defined as a ledge that forms a single level of operation above

which mineral or waste materials are mined back to a bench face. The mineral or waste

is removed in successive layers, each of which is a bench. Several benches may be in

operation simultaneously in different parts of, and at different elevations in the open pit

mine.

Bench elements:

1. Toe,

2. Crest,

3. Floor,

4. High wall (face),

5. Slope angle (angle of inclination) :α,

6. Bench height,

7. Working area,

8. Safety berm.

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Bench parameters

1. Bench height:

The bench height is the vertical distance between each horizontal level of the pit.

The height will depend on :

1. The physical characteristics of the deposit,

2. The degree of selectivity required in separating the ore

3. The size and type of equipment to meet the production requirements,

4. Climatic conditions.

The bench height should be set as high as possible within the limits of the size and

type of equipment selected for the desired production.

The bench should not be so high that it will present safety problems of towering

banks of blasted or un-blasted material.

The bench height in open pit mines will normally range from 15m in large copper

mines to as little as 1 m in uranium mines.

So bench height = 9 m, 3m for every left

2. Bench slope angle:

It's the angle between High wall (face) and the horizontal plane and for granite "hard

rock it's hardness about 7" bench slope angle varies from 70 º to 90 º

α = 5

Because of:

1. Relative small bench height,

2. Hard rock.

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3. Working area:

The width of working bench areas varies within limits depending upon the

equipment used and the mode of operation.

— Wb = 1.2C + Ws + Ws.b + L

Wb : working bench width

C : cut width

Ws : trucks movement "10-15"

L : marginal area "10-20m―

Ws.b = 2 h » h= safty berm height

h= The height of the berm should be of the order of the tire rolling radius,

h=1.5m

Wb = 1.2*18 + 5 + 3 + 5

Wb = 35 m

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4. Bench number:

Bench number = pit averall height / bench height

=295/9

=33 benches

Bench parameters

Bench parameter value

Bench height 9 m

Bench width 35 m

Bench slope angle 5

Number of benches 33

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FIG 7: Cut sequence for sequential pushbacks

Layout of excavating bench

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The time sequence showing shovel loading with single spotting.

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Main haul road design

Over the past three decades, off-highway trucks have been developed for surface

mines up to 400 tons (360 t) in capacity, although 200 tons (180 t) is generally the

largest size in fleet operation. They represent a huge capital investment and a

significant percentage of total costs (Table 13.4.1). If haul road design is inadequate,

the trucks can be highly lethal in the confines of a surface mine. Despite these facts,

haul road design until recently has received little attention.

Haul road design factors must ensure:

1. Minimum costs on a net present value basis for the transport of mineral and

waste throughout the life of the mine.

2. A minimum of traffic congestion and the maintenance of safe, ready access to

the mining operations.

3. The avoidance of areas where slope stability problems could occur.

4. The uses of long-life haul roads rather than short-life roads. This reduces haul

road overall construction costs and operating costs as well as reducing the

demand for haul road construction materials which may not be available in

sufficient quantities from the overburden.

Other factors include the locations of mineral preparation plants, stock yards,

external waste dumps, environmental constraints, etc. All these factors direct attention

to:

1. Haul road layout.

2. Haul road geometry.

3. Haul road construction materials.

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Haul road layout:

For opening up through deposit may layouts of haul road available, like:

1. Switch back road:

They are used to gain altitude or depth without sharp turns.

Fig 8 switch back road layout.

2. Loop road way:

This method is ordinarily employed in mining of deep seated deposits with the use

of motor transport in inclined ingoing trench is driven from the ground surface along

the non-productive side of

open pit at a maximum

permissible gradient at the

end of the open pit field the

road line is turned by 180º.

Fig 8 loop route layout

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3. Spiral route:

The basic conditions for the use of this

method are a level or slightly inclined altitude

of the deposit circular or aver out lines and

considerable size of deposit.

For uphill and downhill pit spiral route

layout used for reaching ore because of relative

small pit geometry.

Haul road geometry:

Number of Lanes:

In-pit roads are usually constructed for single-lane, unidirectional traffic or

two-lane, directional traffic

1. Because traffic density may not be high or

2. Because of space problems. Haul roads from the pit to external waste dumps,

preparation plants, etc.

Number of lanes = 2

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Haul road width:

WR = n (wvehicle+1) +1

Wvehicle: Vehicle width

n: Number of lanes

WR = 2 (4 + 1) +1

Haul road width=11 m

Gradients:

Maximum gradients may be statutorily limited to between 8 to 15% (5 to 8.5°) for

sustained gradients, but in general when considering the economics of uphill haulage,

as well as downhill safety, the optimum gradient for most situations is about 8% (4.5°)

but up to 12% (6.8°) for trolley-assist trucks. For safety and drainage reasons, long

steep gradients should include 150ft (50-m) long sections with a maximum gradient of

2% (1°) for every 1500 to 1800 ft. (500 to 600 m) of severe gradient.

Gradient = 8% (4.5°) to save power and truck maintenance

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Main haul road length

Length of main haul road from pit to waste dump and to processing plant = 2 km.

Reference:

1.SME mining engineering handbook

2. introductory to mining engineering

3. Prof.DR\ Mohamed Abd EL Tawab el gendy

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Drilling and balsting

PART I

Drilling principles

1.1 Introduction:

In virtually all forms of mining, rock is broken through drilling and

blasting. Except in dimension stone quarrying, drilling and blasting are

required in most surface mining. Only the weakest rock, if loosely

consolidated or weathered, can be broken without explosives, using

mechanical excavators (ripper, wheel excavators, shovels etc.) or

occasionally a more novel device, such as a hydraulic jet. In the mining

cycle, drilling performed for the placement of explosives is termed

production drilling. Drilling is also used in surface mining for purposes

other than providing blast-holes. There are minor applications of rock

penetration in surface mining other than drilling. In quarrying,

dimension stone is freed by cutting, channeling, or sawing.

1.2 Classification of methods:

1. Mechanical attack

2. Thermal attack

3. Fluid attack

4. Sonic attack

5. Chemical attack

6. Other methods of attack (electrical, light, or nuclear)

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1.3 PENETRATION RATE:

A function of:

The rock.

The drilling method.

The size & type of bit

The rock properties which effect penetration rate are:

Hardness.

Texture.

Breaking characteristic.

Formation .

1.4 DRILLING METHOD

1.4.1.ROTARY

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1.4.2. ROTARY-PERCUSSION

1.4.3. DOWN HOLE (DH):

The DH drill provides striking energy directly to the bit. There is

rotation so the bit strikes fresh rock with each blow.

Down Hole Drill maintains constant penetration rate at all depths.

Compressed air conducted through the drill steel is used to flush the drill

cuttings from the hole. Performance will not decrease as depth increases.

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PART 2

EXPLOSIVES ENGINEERING

2.1 INTRODUCTION :

The use of explosives in mining and construction applications dates back

to 1627. From 1627 through 1865, the explosive used was black powder.

Black powder was a different type of explosive than the explosives used

today. In 1865, Nobel invented nitroglycerin dynamite in Sweden. He

invented gelatin dynamites in 1866. These new products were more

energetic than black powder and performed differently since

confinement of the explosive was not necessary to produce good results,

as was the case with black powder. From 1867 through the mid-1950's,

dynamite was the workhorse of the explosive industry.

In the mid-1950's' a new product appeared which was called ANFO,

ammonium nitrate and fuel oil. This explosive was more economical to

use than dynamite. During the decades of the 1970's and the 1980's,

ANFO has become the workhorse of the industry and approximately

80% of all explosives used in the United States was ammonium nitrate

and fuel oil.

Other new explosive products appeared on the scene in the 1960's and

1970's. Explosives, which were called slurries or water gels, have

replaced dynamite in many applications. In the late 1970's' a

modification of the water gels called emulsions appeared on the scene.

The emulsions were simple to manufacture and could be used in similar

applications as dynamites and water gels. Commercial explosives fall

into three major generic categories, dynamites, blasting agents and

slurries (commonly called water gels or emulsions).

Blasting problems generally result from poor blast design. Poor

execution in drilling and loading the proposed design and because the

rock mass was improperly evaluated.

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Blast design parameters such as burden, stemming, subdrilling, spacing

and initiation timing must be carefully determined in order to have a

blast function efficiently, safely and within reasonable vibration and air

blast levels.

2.2 SOURCES OF EXPLOSIVE'S ENERGY :

Two basic forms of energy are released when high explosives react. The

first type of energy will be called shock energy. The second type will be

called gas energy.

Figure 2.1 Pressure Profiles for Low and High Explosives

2.2.1 SHOCK ENERGY:

In high explosives, a shock pressure spike at the reaction front travels

through the explosive before the gas energy is released.

The shock energy is commonly believed to result from the detonation

pressure of the explosion. The detonation pressure is a function of the

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explosive density times the explosion detonation velocity squared and is

a form of kinetic energy.

Where:

P = Detonation pressure (Kbar, 1 Kilobar a 14,504 psi)

= Specific gravity of the explosive

= Detonation velocity (fth)

2.2.2 GAS ENERGY:

The gas energy released during the detonation process causes the

majority of rock breakage in rock blasting with charges confined in

boreholes. The gas pressure. often called explosion pressure. is the

pressure that is exerted on the borehole walls by the expanding gases

after the chemical reaction has been completed.

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PART 3

MECHANICS OF ROCK BREAKAGE

3.1 SHOCK ENERGY IN ROCK BREAKAGE :

Unconfined charges placed on boulders and subsequently detonated

produce shock energy which will be transmitted into the boulder at the

point of contact between the charge and the boulder.

3.2 CONFINED CHARGES IN BOREHOLES:

Three basic mechanisms contribute to rock breached with charges

confined in boreholes. The first and least significant mechanism of

breakage is caused by the shock wave.

At most, the shock wave causes microfractures to form on the borehole

walls and initiates microfractures at discontinuities in the burden. This

transient pressure pulse quickly diminishes with distance from the

borehole and since the propagation velocity of the pulse is

approximately 2.5 to 5 times the maximum crack propagation velocity,

the pulse quickly outruns the fracture propagation.

The two major mechanisms of rock breakage results from the sustained

gas pressure in the borehole.

Failure by this mechanism has been recognized for many years and is

commonly called radial cracking (Figure 3.1) .

Figure 2.2 Radial Cracking in Plexiglass

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Direction and extent of the radial crack system can be controlled by the

selection of the proper distance from the borehole to the face (burden)

(Figure 3.2).

Figure 3.2 Influence of Distance to Face on Radial Crack System

The second major breakage mechanism occurs after the radial cracking

has been completed.

3.3 BENCH STIFFNESS :

In most blasting operations, the first visible movement occurs when the

face bows outward near the center. (Figure 2.3).

(Figure 3.4) Axisymmetric Bending Diagram

3.4 EFFECTS OF BLASTHOLE LENGTH :

The rock breakage process occurs in four distinctive steps. As the

explosives detonates. A stress wave moves through the rock uniformly

in all directions around the charge. Radial cracks then propagate

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predominantly toward the free face. After the radial cracking process is

finished. High pressure gases penetrate into the cracks approximately

2/3 of the distance from the hole to the face throughout the radial crack

system. Only after the gas has time to penetrate into the crack systems

are the stresses on the face of sufficient magnitude. To cause the face to

move outward.

The first was to determine the effect of the bench height on the bending

and flexural failure, and the second was to determine the effect of

changing geologic conditions on the movement of the burden itself

(Figure 3.5).

Figure 3.5 Finite Element Model Configurations

3.5 BLASTING PARAMETERS :

In order to compare the model's behavior with that of actual field results,

parameters were chosen so that actual burden movement could be

predicted. The model consisted of a single hole, four inches in diameter

(Figure 3.6).

Figure 3.6 Free Body Diagram for Simulated Condition of Bench

Blasting

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For discussion purposes, four different L/B ratios, 1.2, 2.4, 3.6, and 4.0

will be cosidred .

Figure 3.7 XZ-View of the Deformed Geometry Configuration as LIB

Ratio Changes from 1.2 to 4.0

A close correlation between the finite element model and the rectangular

cross section was observed as hown in Figure 3.9.

Figure 3.9

3.6 GEOLOGICAL EFFECTS ON DISPLACEMENT :

In order to analyze the significance of beds of different materials on

bench blasting, In order to analyze the significance of beds of different

materials on bench blasting, five different models were analyzed using

the same finite element model (Figure 3.11).

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Figure 3.11 Geologic Structure of Models

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PART 4

EXPLOSIVE PRODUCTS

4.1 ENVIRONMENTAL CHARACTERISTICS OF EXPLOSIVES :

The selection of the type of explosive to be used for a particular task is

based on two primary criteria. The explosive must be able to function

safely and reliably under the environmental conditions of the proposed

use, and the explosive must be the most economical to use to produce the

desired end result.

4.1.1 SENSITIVENESS :.

4.1.2 WATER RESISTANCE :

4.1.3 FUMES :

4.1.4 FLAMMABILITY :

4.1.5 TEMPERATURE RESISTANCE :

4.1.6 COLD RESISTANCE :

4.1.7 VELOCITY :

4.1.8 DETONATION PRESSURE :

4.1.9 DENSITY :

Where:

De = Diameter of explosive (in)

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4.1.10 STRENGTH :

4.1.11 COHESIVENESS :

4.2 COMMERCIAL EXPLOSWES :

The products used as the main borehole charge can be broken into three

generic categories. Dynamite, slurries and blasting agents.

4.2.1 DYNAMITE :

Dynamites are the most sensitive of all the generic classes of explosives

used today.

4.2.2 GELATIN DYNAMITE :

Gelatin dynamite, used in commercial applications, can be broken into

three subclasses, straight gelatin, ammonia gelatin and semigelatin

dynamites.

4.2.3 SLURRY EXPLOSIVES

Slurry explosive is a mixture of ammonium nitrate or other nitrates and

fuel sensitizers which can either be a hydrocarbon or hydrocarbons and

aluminum. In some cases explosive sensitizers, such as TNT or

nitrocellulose, along with varying amounts of water are used (Figure

4.5). An emulsion is somewhat different from a water gel or slurry in

characteristics, but the composition contains similar ingredients and

functions similarly in the characteristics, but the composition contains

similar ingredients and functions similarly in the blast hole (Figure 4.6).

In general, emulsions

have a somewhat higher

detonation velocity and

in some cases, may tend

to be wet or adhere to

the blast hole causing

difficulties in bulk

loading. For discussion

purposes, emulsions

and water gels will be treated under the generic family of slurries.

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PART 5

PRIMER AND BOOSTER

5.1 PRIMER AND BOOSTER SELECTION :

The difference between a primer and booster is in its use. Rather than in

its physical composition or makeup. A primer is defined as an explosive

unit which contains an initiator, a booster is used to put additional energy

into a hard or tough rock layer (Figure 5.1).

Figure 5.1 Primer and Booster in Borehole

5.2 SELECTION CRITERIA FOR PRIMER :

The two most critical criteria in primer selection are primer composition

and primer size.

5.3 BOOSTER :

Boosters are used to intensify the explosive reaction at a particular

location within the explosive column. In general, boosters are used to put

more energy into a hard layer within the rock column.

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PART 6

BLAST DESIGN

The design of blasts must encompass the fundamental concepts of ideal

blast design, which are then modified when necessary to account for

local geologic conditions. In order to evaluate a blasting plan, the plan

must be taken apart and each variable or dimension must be evaluated. A

plan must be designed and checked one step at a time. This chapter will

lay out a step-by-step procedure for the analysis of a blasting plan.

Methods to determine whether design variables are in normally

acceptable ranges will be examined.

6.1 BURDEN :

Burden distance is defined as the shortest distance to relief at the time the

hole detonates (Figure 6.1).

The selection of the proper burden is one of the most important decisions

made in any blast design. Of all the design dimensions in blasting. It is

the most critical.

Figure 6.1 Symbols for Blast Design

Where:

B = Burden

T = Stemming

J = Subdrilling

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L = Bench height

H = Blasthole depth

PC = Powder column length

If the operator has selected a burden and used it successfully for a drill

hole of another size and wants to determine a burden for a drill hole that

is either larger or smaller, one can do so quite easily if the only thing that

he is changing is the size of the hole and the rock type and explosives are

staying the same. To do this, one can use the following simple ratio:

Where:

B1 = Burden successfully used on previous blasts

De1 = Diameter of explosive for B1

B2 = New burden

De2 = New diameter of explosive for B2

6.1.1 ADJUSTMENTS FOR ROCK & EXPLOSIVE TYPE :

When an operator is moving into a new area where he has no previous

experience, he would have only general rock and explosive

characteristics to work with. When moving into a new area, especially

one where there are residents nearby, it is essential that the first shot not

be a disaster. To approximate burden under these situations, the

following empirical formula is helpful.

[

]

Where:

B = Burden (ft)

SGe = Specific gravity of explosive

SGr = Specific gravity of rock

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The previous equations proposed for burden selection used the specific

gravity of the explosives as an indicator of energy. The new generation

of slurry explosives called emulsions have somewhat different energies

but near constant specific gravity. The equation that uses relative energy

is:

√

Where:

B = Burden (ft)


Stv = Relative bulk strength (ANFO = 100)

SGr = Specific gravity of the rock

6.1.2 GEOLOGIC CORRECTION FACTORS :

No one number will suffice as the exact burden in a particular rock type

because of the variable nature of geologyTo estimate the deviation from

the normal burden formula for unusual rock structure, two constants are

incorporated into the formula. Kd is a correction for the rock deposition

and Ks is a correction for the geologic structure. Kd values range from

1.0 to 1.18 and describe the dipping of the beds (Table 6.3).

TABLE 6.3 CORRECTIONS FOR ROCK DEPOSITION

BEDDING ORIENTATION Kd

Bedding steeply dipping into cut 1.18

Bedding steeply dipping into face 0.95

Other cases of deposition 1.00

The correction for the geologic structure takes into account the fractured

nature of the rock in place, the joint strength and frequency as well as

cementation between layers of rock. The correction factors for rock

structure ranges from 0.95 to 1.30 (Table 6.4). Massive intact rock

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would have a Ks value of 0.95 while heavily broken fractured rock could

have a Ks value of about 1.3.

TABLE 6.4 CORRECTIONS FOR GEOLOGIC STRUCTURE

GEOLOGIC STRUCTURE Ks

Heavily cracked, frequent weak joints weakly cemented

layers

1.30

Thin well-cemented layers with tight joints 1.10

Massive intact rock 0.95

6.2 STEMMING :

The common material used for stemming is drill cuttings, since they are

conveniently located at the collar of the blast hole.

T = 0.7 x B (For crushed stone or drilling chips)

Where:

T = Stemming (ft)

B = Burden (ft)

Figure 6.2 Stemming Zone Performance

6.3 SUBDRILLING :

Subdrilling is a common term to denote the depth which a blasthole will

be drilled below the proposed grade to ensure that breakage will occur to

the grade line. Blastholes normally do not break to full depth.

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J= 0.3×B

Where:

J = Subdrilling (ft)

B = Burden (ft)

Figure 6.5 Problems of Soft Seam off Bottom

5.4 BLASTING CONSIDERATIONS :

The blasting consideration of fragmentation, air blast, flyrock and

ground vibration would have to be assessed.

The more massive the rock in a production blast. The more probable the

outcome listed in Table 5.5.

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TABLE 5.5 POTENTIAL PROBLEMS AS RELATED TO

STIFFNESS RATIO (L/B)

6.5 SPACING :

If holes are initiated simultaneously. Spacing's must be spread further

apart than if holes are timed on a delay. (Figure 6.6).

Figure 6.6 Shattered Zones from Close Spacing

Figure 6.7 Rough Walls from Excessive Spacing

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6.5.1 INSTANTANEOUS INITIATION LOW BENCHES

In order to check the blasting plan and determine if spacing is within

normal limits, the following equation can be used:

Where:

S= Spacing (ft)

L = Bench height (ft)

B = Burden (ft)

6.5.2 INSTANTANEOUS INITIATION HIGH BENCHES

6.5.3 DELAYED INITIATION LOW BENCHES :

When the stiffness ratio is between one and four with delayed initiation

between holes, the following relationship is used to check spacing:

Where:

S = Spacing (ft)

L = Bench height (ft)

B = Burden (ft)

6.5.4 DELAYED INITIATION HIGH BENCHES :

S=1.4 B

Where:

S = Spacing (ft)

B = Burden (ft)

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If the calculated spacing value is within plus or minus 15% of the actual

spacing, the spacing is within reasonable limits.

6.6 Drilling pattern:

Figure 6.12 V-Cut (Square Corner), Progressive Delays, S = 1.4 B

Figure 6.12 Stagger pattern

6.7 Initiation pattern:

There are two types of initiation patterns

6.7.1 Chevron initiation pattern (closed)

6.7.2 Corner initiation

pattern (flat)

Corner initiation pattern is

used because it gives good

control of rock

displacement and gives

good fragmentation

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PART 7

CALCAULATION

7.1 BURDEN :

√

We take

√

=15.15ˋ

=4.55 m

=4.55×1×0.95

=4.5 m

7.2 STEMMING :

=0.7×4.5

=3.15 m

7.3 SUBDRILLING :

=0.3×4.5

=1.35 m

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7.4 STIFFNESS RATIO :

We take L= 9 m

= 2

7.5 SPACING :

Delayed initiation with the S.R greater than 1 but less than 4

= 5 m

7.6 CHARGE LENGTH :

=7.2 m

7.7The total weight of explosive per Colum :

› loading density

=132.3 kg

7.8VOLUME UNDER AREA :

= 202.5 m

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7.9 POWDER FACTOR :

⁄

7.10Weight of rock broken per hole :

= 570 t

The product of ore per day is 10000 t.

The stripping ratio is 1:5.

The product of broken rock per day is 60000 t.

7.11 The number of hole:

= 105 Hole

We use two face.

For one face length is 65 m and width is 18 m.

For one face 4 rows and 13 Colum.

7.12 THE NUMBER OF DRILLS:

Rate of drilling is 35 m/hour.

Total meters drilled per shift /drill =35 x 5 =175 m/shift

No. of drills

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We use 3 drills

We work 2 shifts per day (2 shift/day)

7.13 The COST OF DRILLING AND BLASTING:

The cost of one cubic meter of drilling and blasting is about 6:10 LE

(We take the cost is ( 8 ⁄

Cost

7.14 CHAPTER 7 SUMMARY

type of explosive is emulsion

1 BURDEN (B) M 4.5

2 STEMMING (T) M 3.15

3 SUBDRLLING (J) M 1.35

4 STIFFNESS RATIO (S.R) 2

5 SPACING (S) M 5

6 CHARGE LENGTH ( ) M 7.2

7 CHARGE PER HOLE ( ) kg 132.3

8 POWDER FACTOR (P.F)

⁄

0.65

9 NUMBER OF HOLE (n) 105

10 ORE PRODUCTION T 10000

11 TOTAL CHARGE ( ) kg 13891

12 NUMBER OF DRILLS 3

13 COST (LE) 168480

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References:

Blasters‘ Handbook, E. I. du Pont de Nemours & Co., Pont de

Nemours & Co., Wilmington, DE W

Rock Blasting and Overbreak Overbreak Control, National

Highway Control, National Highway Institute, Washington, DC

Institute, Washington, DC

―Introductory mining engineering ―, Howard l. Hartman &

JanM.mutmansky, Scound Edition ,new York .2002.

Lectures by Prof. Dr. Mohamed Abd El Tawab El Gendy

Lectures by DR. Ibrahim assakkaf

Lectures by eng. Abd-Elmoneam seleim

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Loading and transportation

Introduction

Equipment selection is one of the most important steps of open pit design. Mining

costs are mainly affected by the number and capacity of equipment.

Equipment selection effects economic considerations in open-pit design,

especially overburden or (waste rock) and ore mining costs and cost increasing

parameters as a function of plan location and depth. Mining costs are a function of site

conditions, operating scale and equipment. The purpose of equipment selection is to

select optimum equipment with minimum cost.

Equipment selection for open-pit mines is a very important decision which will

impact greatly the economics operations.

Selection of loading equipment

The three main types of surface loading equipment:

1. Cable shovels,

2. Hydraulic shovels,

3. Front end loaders.

The following table shows comparison between the Operating

characteristics of each type:

Cable Shovel Hydraulic Shovel Front-end Loader Proven and reliable, low

operating costs and machine

availability.

Significant disadvantages in

un-blasted toes.

No selectivity; high blending

vertically.

Cannot dig below floor level.

Cannot traverse gradients

greater than 1 in 20.

No control over dump rate.

Has long life (20-30 years).

Low operating costs

compared to hydraulic

shovels or FELs.

Powerful digging force permits

the machine to operate in a variety

of conditions.

Half the amount of hoist and

crowd needed for same size

bucket as a cable shovel.

Can be used as back hoe and can

dig below the floor.

Half the size of shovel for same

bucket size as cable shovels.

Double speed of cable shovel.

Very high selectivity.

Horizontal crowd motion can load

thin seams.

Discharge into trucks can be

controlled.

Versatile, highly mobile,

multipurpose.

Increasingly used as LHD machines.

Traction and stability problems.

Bucket wider than front tires (tire

protection).

Tire costs considerable.

Not suitable for soft grade. High

maneuvering required, increasing risk

of overturning or falling off edge.

Well suited for: well fragmented

ground on good under-footing, road

maintenance, stockpile reclamation

and flexible back-up machine.

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Characteristics of local operating condition in the proposed mine:

badly fragmented product (R.O.M) on bad under-footing.

Front-end Loader is no longer useful.

High selectivity is needed to load the ore and the waste separately.

Electric Cable Shovel is no longer useful.

Thus,for more selectivity we have to use proper sized Hydraulic Shovel as back

hoe

Selection of haulage equipment

The three main types of surface haulage equipment which are generally

available for mine use are:

1. Belt Conveyor Haulage.

2. Locomotive Haulage.

3. Truck Haulage.

The following table shows comparison between the Operating characteristics

of each type:

Equipment Belt Conveyor Haulage Locomotive Haulage Truck Haulage

Operation(output) Continuous (not cyclic) Cyclic (discontinuous) Cyclic (discontinuous)

Capacity (t/hrs.) High high Moderate

Haulage distance

(km)

Limited (up to 5 km) unlimited Limited (0.5 – 10 km)

Relief Moderate low Moderate

Grade High (Up to 20o) Low (up to 3

o) Moderate (8

o – 12

o)

Advantages High output at relative

low cost.

Better grade ability

High daily output.

Long haulage distance.

Coarse / blocky product.

High flexibility.

High maneuverability.

Moderate grade

ability.

Coarse product.

Disadvantages Low flexibility.

High installation cost.

Limited to fine product.

Poor grade ability.

High installation cost.

High operating cost

(esp. with bad roads)

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Characteristics of local operating condition in the proposed mine:

The grade is moderate.

Locomotive Haulage must be excluded.

The daily Capacity is moderate about (10,000 t/sh.), haulage distance is

moderate (2 km), high flexibility and maneuverability is needed for advanced

face.

Belt Conveyor Haulage must be excluded.

Thus, for more flexibility and maneuverability we have to use Truck Haulage.

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Estimation of proper size and number of Loading and haulage

equipment

Design and operating data:

Ore: gold with a cut of grade equal 0.5 g/t

Associated rock is granite, well blasted

Required production of ore = 10000 t/day

Waste associated with ore, 50000 t/day (source from mine field)

Total required production = 60000 t/day

Two faces (two shovels) are selected to cover the required production

Working time:3 shift/day, 8 hours / shift, effective time of loading and

haulage 6 hours/shift

Haulage distance to crushing plant = 2000 m

Haulage distance to stock pile = 2000 m

Job condition and management is considered to be good

Loading equipment selection

Proposed production per face = 60000/2=30000 t/day (One face)

Proposed production per shift = 30000/3=10000 t/sh. (One face)

Cycle time range of loader = 28- 45 sec,

We consider as an average cycle time=35 sec

SO No. of cycles per hour (C) = 3600/35 = 102cycles/hr. and the actual

bucket size used calculated by:

BC = [Q (m3 bank/ hr) * swell factor] / [C * S * O.A * BF]

Where:

(BC)=bucket capacity (m3)

(Q)= productivity (m3 bank/hr.)

(C)= no. of cycles per hours [100 cycle/hr.]

(S)=swing factor

(O.A)=operating efficiency

(BF)=bucket fill factor

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From table [13.3.3 SME handbook] @ angle of swing = 60 o

Swing factor (s) = 1.1

From table [13.3.4 SME handbook] @ good job and good

management conditions

Operating efficiency (O.A) =0.73

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From table [13.3.5 SME handbook] @ granite material type

Bucket fill factor (fillability) (BF) = 0.8

Bank density (ρbank) =2.41 t/ m3

Swell factor = 1.55

Loose density (ρloose)=2.41/1.55=1.55 t/m3

Shovel production per hour (t/hr)=10000/6=1666.67 t/hr

Shovel production per hour (m3 bank /hr) = (t/hr)/ ρbank=1666.67 /2.41=691.56 m

3 bank/hr

BC= [Q (m3 bank/ hr) * swell factor] / [C * S * O.A * Bf]

BC = (691.65 * 1.55) / (102 * 1.1 * 0.73 * 0.8) = 16.36 m3

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The nearest mining excavator size of bucket (17 m3), we will use one back

hoe (for more selectivity) hydraulic excavator RH 120 E (TEREX) with

bucket size (17m3)

Actual productivity of loader = BC * C * S * O.A * Bf * ρloose =17 * 102*

1.1*0.73*0.8*1.55= 1726.57 t/hr

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Actual productivity of loader (t/sh.) =1726.27*6=10359.47 t/sh.

(Satisfy face shift production)

Daily production of one face = 10359.47*3=31078.4 t/day

Total daily production of the two faces =31078.4*2=62156.8 t/day

Haulage equipment selection;

Bucket capacity per one pass = BC * BF * ρloose

=17*0.8*1.55

=21.08 t/pass

Where:

BC=bucket capacity (m3),

BF=bucket fill factor,

ρloose= the loose density of material.

To fill the truck we use 4:6 passes:

In case of 4 passes:

Truck capacity=21.08*4=84.32 t

Difference (%) = (90.9-84.32)/90.9=7.2%


Truck capacity=21.08*5= 105.4t

Difference (%) = (109-105.4)/109=3.3%


Truck capacity=21.08*6=126.48 t

Difference (%) = (136-126.48)/136=7%

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So we use truck capacity of (109) [MT 3000] and fill it with (5) passes

Truck cycle time = loading time + haulage time (loaded) + dumping time +

returning time (empty) +spotting time

From [table 9.3.4 SME handbook] @ average conditions and Rear-dump truck

type

Dumping time =1.3 min

From [table 9.3.5 SME handbook] @ average condition and Rear-dump truck

type

Spotting time = 0.3 min

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Loading time = no. of passes * loader cycle time =5*35=175 sec=3 min

Haulage distance (Loaded) is about 2000 m is divided into :-

200 m face @ speed = 15 km/hr

200 m ramp @ speed = 25 km/hr

1400 m main road @ speed = 40 km/hr

200 m dump area @ speed = 15 km/hr

Returning distance (Empty) is about 2000 m is divided into :-

200 m face @ speed =20 km/hr

200 m ramp @ speed = 30 km/hr

1400 m main road @ speed = 45 km/hr

200 m dump area @ speed = 20 km/hr

Hauling time (loaded) =

200/ (1000*15) +200/ (1000*25) +1400/ (1000*40) +200/ (1000*15) = 0.0697

hr=4.182 min

Returning time (empty) =

200/ (1000*20) +200/ (1000*30) +1400/ (1000*45) +200(1000*20)

=0.058hr=3.47 min

Truck cycle time = 3+4.182+1.3+3.47+0.3=12.25 min

No. of truck cycles per shift = (60*6)/12.25= 29 cycle/shift

Productivity of one truck /sh. =105.4*29=3056.6 t/sh.

Req. number of trucks= (loader productivity /sh.)/ (one truck

productivity/sh.)=10359.47/3056.6= 4 trucks

Actual required no. of truck =4/0.83=5 trucks

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Synchronization of loader and trucks

n ( tspotting + tloading ) >= tcycle

4(0.3+3)>=12.25

13.2>12.25

Difference = 13.2-12.25=0.95 min

Reference

1. SME mining engineering hand book

2. Introductory to mining engineering

3. Prof.Dr\ Mohamed abd el tawab el gendy's lectures

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Ore dressing

2.1.Crushing

Crushing is the first mechanical stage in the process of comminution in which the

main objective is the liberation of the valuable minerals from the gangue.

2.2 .Primary crushers

Primary crushers are heavy-duty machines, used to reduce the run-of-mine ore down

to a size suitable for transport and for feeding the secondary crushers or AG/SAG

mills. They are always operated in open circuit, with or without heavy-duty scalping

screens (grizzlies). There are two main types of primary crusher in metalliferous

operations –jaw and gyratory crushers- although the impact crusher has limited use

as a primary crusher and will be considered separately.

2.3 .Jaw Crusher

Jaw Crusher is one of the main types of primary crushers in a mine or ore processing

plant. The size of a jaw crusher is designated by the rectangular or square opening at

the top of the jaws (feed opening). For instance, a 24 x 36 jaw crusher has a opening

of 24" by 36", a 56 x 56 jaw crusher has a opening of 56" square. Primary jaw

crushers are typically of the square opening design, and secondary jaw crushers are

of the rectangular opening design. However, there are many exceptions to this

general rule. A Jaw Crusher reduces large size rocks or ore by placing the rock into

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compression. A fixed jaw, mounted in a "V" alignment is the stationary breaking

surface, while the movable jaw exerts force on the rock by forcing it against the

stationary plate. The space at the bottom of the "V" aligned jaw plates is the crusher

product size gap, or the size of the crushed product from the jaw crusher. The rock

remains in the jaws until it is small enough to pass through the gap at the bottom of

the jaws

2.4 .Grinding

Grinding is a powdering or pulverizing process using the rock mechanical forces of

impaction, compression, shearing and attrition.

Milling, sometimes also known as fine grinding, pulverising or comminution, is the

process of reducing materials to a powder of fine or very fine size. It is distinct from

crushing or granulation, which involves size reduction to a rock, pebble or grain size.

Milling is used to produce a variety of materials which either have end uses

themselves or are raw materials or additives used in the manufacture of other

products.

A wide range of mills has been developed for particular applications. Some types of

mills can be used to grind a large variety of materials whereas others are used for

certain specific grinding requirements. This brief aims to present the factors to

consider when choosing a particular grinding applications and to give an overview of

the equipment which is available.

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The two main purposes for a grinding process are:

• To liberate individual minerals trapped in rock crystals (ores) and thereby

open up for a subsequent enrichment in the form of separation.

• To produce fines (or filler) from mineral fractions by increasing the specific

surface.

2.5. Ball Mill

Ball mills are similar in concept to the rod mill but are charged with steel balls in

place of the rods. The mill consists of a cylindrical drum, sometimes tapered at one

end, and usually has a charge of steel balls (up to 40% by volume) ranging in size up

to

125mm for larger mills.

Product size can be as

small as 0.005mm, but

product size is dependant

upon the time the charge

spends in the grinding zone

and therefore the reduction

rate is a function of the throughput. The lining material is of great importance as

there is a significant amount of wear taking place due to the action of the steel balls.

The speed of rotation is optimum at about 75% of critical speed. Some mills are

compartmentalized with each subsequent section having a smaller ball size. The

mineral can pass through to the proceeding section, but the balls cannot. This ensures

that the smaller particles are attacked by the smaller grinding media. It is a versatile

grinding mill and has a wide range of applications. The mill can vary in size from

small batch mills up to mills with outputs of hundreds of tonnes per hour. They are

the most widely used of all mills. Small hand operated ball mills are used in Bolivia

for preparation of ore, sand and

gravel.

3.1. Ore characteristics

Ore competency and hardness is

the prime determinant of the circuit

configuration. Low competency

may permit the use of lower capital

cost crushing and milling

equipment such as MMD sizers,

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single-stage primary mills and SAG mills. High competency will dictate the

examination of SABC circuits and, for ores that exhibit extreme resistance to

breakage, staged crushing (two or three stages) followed by either ball or SAG

milling may have to be considered. At higher throughputs, such circuits can consider

the use of HPGRs in place of SAG milling. Circuits handling moderately competent

ores can take advantage of operating cost savings by employing fully autogenous

grinding (AG) mills.

3.1.Free-Milling Ore Process Options

The recovery circuits of choice are either carbon-in-leach (CIL) or heap leach

followed by carbon-in-solution (CIS) . For free milling ores exhibiting a high gold

recovery at a reasonably coarse grind size and with average oxygen and cyanide

consumptions, the engineer is faced with few selection issues in determining the

flowsheet.

3.2. Gravity-recoverable gold

The benefits of gravity recovery can be readily assessed by undertaking leach tests

with and without pre-treatment coupled with mineralogical analysis and examination

of gold leach tails solids. Early equipment included shaking tables, spirals, drums

and other devices. The alternatives expanded to more sophisticated and efficient

centrifugal separators such as Knelson and Falcon concentrators with the latter

perhaps being more applicable to the treatment of finer solids. In recent times, a

further range of gravity equipment has been successfully commercialized including

in-line pressure jigs. equipment development has also extended to the use of

high-intensity cyanidation devices to solubilize gold from the concentrates.

3.3. Complex Ore Process Options

Complex ores are intermediate between free-milling and refractory ores. As such,

they give rise to high usages of cyanide and oxygen and/or are pregrobbing.

Speciation of cyanide complexes within leach liquors especially for feeds known to

contain copper, zinc, thiosulfates and other complexes is recommended. Flowsheet

selection issues are discussed under the following headings.

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4.1. Methods of treatment gold ore:

1. Gravity concentration

2. leaching

3. Flotation

4.1.1.Gravity concentration

Gravity separation, one of the oldest separation techniques, has become increasingly

popular in modern plants, with new equipment enhancing the range of separations

possible (Laplante and Doucet, 1996). When coupled with generally low capital and

operating costs and lack of chemicals to cause environmental concerns, this often

provides an attractive process for the recovery of gold. Gravity separation relies

upon the differences in density of minerals to provide efficient separation. The ease

and efficiency of separation is dependent on a number of factors, including relative

density, particle size and shape, liberation – all of which affect the selection of

equipment type.

In the case of gold, gravity tools can be useful in solving a number of problems.

These can include what is termed spotty or coarse gold, which makes mass balancing

and gold accounting extremely difficult. By utilizing gravity ahead of the leach train,

early recovery of gold in the process can also have financial benefits and avoid

potential losses. Gravity recovery is also a useful diagnostic tool and can, and has

been, used to check for the potential salting of samples. Removal of coarse

gravity-recoverable gold can also enhance leach kinetics in plant practice. Use of

gravity recovery as a safety net on tailings has also been exploited at several

operations, where unleached gold, either ascoarse particles or sulfide locked, are

recovered from the tailings by gravity

means and re-treated, usually with a re-grind prior to a re-leach. The range of

equipment available for gravity separation includes standard mineral jigs, Kelsey

jigs, In-Line Pressure jigs, spirals, tables, Mozeley sizer, Knelson, Falcon Superbowl

and others.

1.1. Conventional jigs

Conventional jigs are often used to recover heavy minerals that are liberated at a

coarse particle size from crushing/grinding circuits, thus avoiding subsequent

over-grinding and loss.

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1.2. Centrifugal jigs

Centrifugal jigs use enhanced forces generated by their spinning motion to enable

finer particle sizes and closer specific gravity (SG) minerals to be separated. The

Kelsey jig is the most common example of this type of separator.

1.3. Spirals

Spirals are one of the oldest gravity separators. There is a wide range of profiles

available including low-grade, medium-grade, high-grade and fine mineral models,

plus ones incorporating different wash water techniques.

Careful monitoring and control of size distribution is important in achieving

optimum results with spirals.

1.4. Mozley gravity separator (MGS)

The MGS is a low-capacity high-performance gravity separator suitable for treating

difficult fine particle feeds below 75 mm.

1.5. Falcon and Knelson concentrators

These are centrifugal type gravity separators also suited to fine particle-size feeds

(Ancia et al., 1997). These units come in batch and continuous configuration for both

laboratory testing and operational application.

1.6. Shaking tables

Tables are often used in the laboratory as a preliminary test to ascertain an ore‘s

amenability to gravity separation or upgrade, or as a tool in their own right. Size of

tables used in the laboratory environment vary but usually range from third or

quarter production size up to half and on occasion full size.

1.7. Super-panners

A laboratory diagnostic tool used to produce the highest concentration of material,

super-panners are often used in conjunction with other gravity devices as a final

cleaning step, and in the preparation of samples for mineralogical work.

4.2.Leaching

Oxides are leached with a sulfuric acid or sodium carbonate solvent, while sulfates

can be leached with water or sulfuric acid. Ammonium hydroxide is used for native

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ores, carbonates, and sulfides, and sodium hydroxide is used for oxides. Cyanide

solutions are a solvent for the precious metals, while a sodium chloride solution

dissolves some chlorides. In all cases the leach solvent should be cheap and

available, strong, and preferably selective for the values present.

Leaching is carried out by two main methods: simple leaching at ambient

temperature and atmospheric pressure; and pressure leaching, in which pressure and

temperature are increased in order to accelerate the operation. The method chosen

depends on the grade of the feed material, with richer feed accommodating a costlier,

more extensive treatment.

4.2.1.Type of leaching :

1-Leaching in-place, or in situ leaching, is practiced on ores that are too far

underground and of too low a grade for surface treatment. A leach solution is

circulated down through a fractured ore body to dissolve the values and is then

pumped to the surface, where the values are precipitated.

2- Heap leaching is done on ores of semilow grade--that is, high enough to be

brought to the surface for treatment. This method is increasing in popularity as larger

tonnages of semilow-grade ore are mined. The ore is piled in heaps on pads and

sprayed with leach solution, which trickles down through the heaps while dissolving

the values. The pregnant solution is drained away and taken to precipitation tanks.

3-Higher-grade ores are treated by tank leaching, which is carried out in two ways.

One method is of very large scale, with several thousand tons of ore treated at a time

in large concrete tanks with a circulating solution. In the second method, small

amounts of finely ground high-grade ore are agitated in tanks by air or by mechanical

impellers. Both solutions pass to precipitation after leaching is completed.

Pressure leaching shortens the treatment time by improving the solubility of solids

that dissolve only very slowly at atmospheric pressure. For this process autoclaves

are used, in both vertical and horizontal styles. After leaching, the pregnant solution

is separated from the insoluble residue and sent to precipitation.

4.2.2.What is heap leaching?

To those of us in the gold industry, the question ‗‗What is Heap Leaching?‘‘ seems to

have an obvious answer. In the simplistic sense, heap leaching involves stacking of

metal-bearing ore into a heap on an impermeable pad, irrigating the ore for an

extended period of time (weeks, months or years) with a chemical solution to

dissolve the sought-after metals, and collecting the leachant (pregnant solution) as it

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percolates out from the base of the heap. Fig. 2 is an aerial photograph showing the

typical elements of a preciousmetals heap-leach operation – an open-pit mine, a heap

of crushed ore stacked on a plastic pad, ponds, a solution process facility for

recovering gold and silver from the pregnant solution, and an office facility. For a

small operation such as the one illustrated here, very limited infrastructure is

required. In a more complex sense, heap leaching should be considered as a form of

milling. It requires a non-trivial expenditure of capital, and a selection of operating

methods that trade off cost versus marginal recovery. Success is measured by the

degree to which target levels and rates of recovery are achieved. This distinguishes

heap leaching from dump leaching. In dump leaching, ores are stacked and leached

in the most economical way possible, and success is achieved with any level of net

positive cash flow. The bibliography of precious metals heap-leaching is quite

extensive, but a limited bibliography has been compiled.

4.2.3. Why select heap leaching as the processing method?

Gold and silver can be recovered from their ores by a variety of methods, including

gravity concentration, flotation and agitated tank leaching. Methods similar to heap

leaching can be employed: dump leaching and vat leaching (vat leaching is the

treatment of sand or crushed ore in bedded vats with rapid solution percolation).

Typically, heap leaching is chosen for basic financial reasons – for a given situation,

it represents the best, or safest, return on investment. Some of the financial

considerations that might result in the selection of heap leaching are presented

below.

4.2.7. Heap leaching of gold and silver ores

Heap leaching of gold and silver ores is conducted at approximately 120 mines

worldwide. Heap leaching is one of several alternative process methods for treating

precious-metal ores, and is selected primarily to take advantage of its low capital cost

relative to other methods. Based on surveys of about 60% of the known heap-leach

operations, it is likely that heap leaching produces 12% of the world‘s gold. In 2004,

at least 10 major heap leaches were in the late design stages, in such diverse locations

as Brazil, Kazakhstan, Laos, Mexico, Peru, the United States and Uzbekistan. Heap

leaching for silver is conducted using the same principles and operating practices as

for gold, but heap-leach operations produce only a small fraction of world silver

production.

Heap leaching had already become a fairly sophisticated practice at least 500 years

ago. illustrates a heap leach with a 40-day leach cycle which could pass in many

ways for a modern heap leach. The Agricola heap-leach recovered aluminum

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(actually alum) for use in the cloth-dying industry. Copper heap and dump leaches in

southern Spain were common by about 1700. Gold and silver heap-leaching began

with the first Cortez heap leach in 1969. While many projects have come and gone,

Cortez is still going – their new 63,000 t/d South Area leach started up in 2002.

The largest U.S. precious-metal heap-leach is the Round Mountain operation in

Nevada, with over 150,000 t/d of ore going to crushed or run-of-mine heaps, at an

average grade of 0.55 g/t (this chapter follows the convention of ton for short ton and

tonne for metric ton; t/d reflects metric tonnes). Worldwide, Newmont‘s Yanacocha

operation in Peru holds the record, with a production rate of 370,000 t/d, at an

average total reserve gold grade of 0.87 g/t. On the other end of the scale, some very

high-grade ores – up to 15 g/t (0.5 oz/ton) – are being successfully processed at rates

of several hundred tonnes per day (Sterling, Nevada; Hassai, Sudan; Ity, Ivory

Coast). Nevada was the ‗birthplace‘ of modern gold heap-leaching in the late 1960s,

and is only now giving up its dominance of this technology. Other very large gold

districts – notably the pre-Cambrian shield areas of Canada, Australia and South

Africa – show relatively few heap leaches. There are several reasons for this

geographic concentration, but the primary reason is that Nevada gold deposits tend

to have been created by low-energy geologic

4.2.8.Chemistry of gold and silver heap leaching

The chemistry of leaching gold and silver from their ores is essentially the same for

both metals, and many ores contain a mixture of the two. A dilute alkaline solution of

sodium cyanide dissolves these metals without dissolving many other ore

components (copper, zinc, mercury and iron are the most common soluble

impurities). Solution is maintained at an alkaline pH of 9.5 to 11. Below a pH value

of 9.5, cyanide consumption is typically high. Above a pH value of 11, metal

recovery decreases.

Silver is usually not as reactive as gold with cyanide. This is because gold almost

always occurs as the metal, whereas silver may be present in the ore in many

different chemical forms, some of which are not cyanide-soluble. Gold recovery

efficiency from operating heap leaches is typically _70%, although it can range from

50 to 90%. Silver recovery efficiency is typically _55%. Other leaching agents, such

as thiosulfate, thiourea, hypochlorite and bromine, have been experimented with an

alternative to cyanide, but cyanide is by far the most effective and the most

environmentally friendly leaching agent.

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4.2.9.Cyanide in Gold Extraction

As part of an overall planning procedure, the role cyanide will play in a mining

operation must be well understood and defined. Personnel who have a sound

knowledge of the types of cyanide and how their chemistry relates to extraction,

recycling, workers' health and environmental risks should participate in planning the

mine and its cyanide strategy.

1. Types and Names of Cyanide Complexes

Because there is such a variety of cyanide complexes, it is often difficult to compare

results of toxicological and environmental investigations of cyanide. For this reason

it is important that people responsible for managing cyanide understand and specify

the type of cyanide being measured. An essential chemical fact about cyanide is

that it is not an element like, for example, gold, arsenic and mercury. The free

cyanide ion comprises a nitrogen atom bonded to a carbon atom (CN-). This ion

combines with hydrogen to form hydrocyanic acid (HCN) and with metal ions to

form salts. The term cyanide is imprecisely applied to all of these forms and more

precise terms are defined as follows:

• Free cyanide-the sum of the free cyanide (CN-) ion and hydrocyanic acid,

HCN(aq). It is the free cyanide ion that is generally measured after suitable sample

treatment;

• Titratable cyanide-the cyanide concentration determined in solution by titration

with silver nitrate (AgNO3); often taken to mean free CN- but may include cyanide

from the dissociation of some cyano-metal complexes;

• simple cyanometal complexes-these contain only one type of metal ion, commonly

an alkaline or alkaline earth metal ion, and dissociate when dissolved in water to

release free cyanide;

• Complex cyanides-these contain more than one type of metal ion and dissociate in

water to release a metal ion and a cyanide-metal ion complex via a reaction of the

type, for example;

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where the complex ion may then dissociate further to give free cyanide;

• Total cyanide-the sum of all of the different forms of cyanide present in a system.

'Total cyanide' is a toxicologically meaningless term since its measurement requires

harsh sample treatment to break down intractable complex cyanides before free

cyanide can be measured;

• Weak acid dissociable (WAD) cyanide-cyanide that is readily released from

cyanide-containing complexes when the pH is lowered. Any free cyanide already

present and cyanide released from nickel, zinc, copper and cadmium complexes (but

not iron or cobalt complexes) is measured. WAD cyanide is measured by treating the

sample with a weak acid buffer solution such as a sodium acetate/acetic acid mixture

at pH 4.5 to 6. This is less harsh than the methods used for total cyanide. WAD

cyanide is generally considered to be the best current measure for assessing human

and animal toxicity;

• Cyanide amenable to chlorination (CATC)-an analytical quantity that requires

similar sample treatment to WAD but is much less reliable; and

• WAD CN, Available Cyanide and CATC generally measure the free and weakly

dissociable cyano-metal complexes

Different forms of cyanide can also be defined by reference to the various analytical

techniques and the types of cyanide that each technique is able to measure. For

example, free cyanide plus weak dissociable cyano-metal complexes may defined as

'WAD CN 4500-CN-I', 'CATC' or 'Available CN Method OIA-1677' depending on

the method used (adapted from Schulz, 2002). This approach to cyanide

nomenclature requires full disclosure of the method and the analytical protocol used

to ensure that meaningful comparisons and check analyses can be undertaken gas

2. Recycling or Disposing of Cyanide

As discussed in Section 4.3, even with the best recycling efforts, there may be

cyanide waste that needs to be treated to enhance its rate of degradation.

4. Natural degradation by oxidation

Reactions that cause the oxidative loss of the cyanide ion from aqueous alkaline

solutions are described by the following equations:

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Reactions 9 and 10 represent oxidative loss of cyanide ion giving rise to cyanate ion

and cyanogen gas respectively. As can be seen from the pH versus the

oxidation-reduction potential stability (Pourbaix) diagram presented in Figure 3, the

cyanate ion is thermodynamically the most stable form of cyanide under the majority

of operating and ambient environmental conditions. However, reaction 9 proceeds

very slowly in the absence of a catalyst so that loss of cyanide by this route is limited

during the ore extraction process.

Equation 8, in which cyanide is degraded by combining in aqueous solutions to

produce ammonia and bicarbonate, is particularly relevant to cyanide analysis and is

discussed further in Section 2.3 below

5. Cyanide: Health & Hygiene In The Workplace

6. Advances in the cyanidation of gold

In the last 15 years, most of the development in the cyanidation of gold has occurred

in response to the decreasing grade of deposits, the shift from surface mining to

underground mining, the increasing complexity of treatment and the concern for

environmental constraints. Research has focused on the optimization of reagent

addition (e.g. cyanide, oxygen and lead nitrate) and on metallurgical strategies to

measure and control these parameters, as well as in the areas of equipment and

automation. Efforts to introduce online cyanide analysers were made in the late

1980s; however, progress was slow.

The analyzers required additional testing to be reliable and effective and to yield

more accurate titration results. Online dissolved oxygen sensors were integrated in

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the late 1980s simultaneously with oxygen addition and various injection devices.

Sulfide in gold ores not only consumes oxygen and cyanide but also forms a coating

on gold grains. This passive layer reduces gold-leaching kinetics and overall

extraction. Improvement of lead nitrate addition strategy has made it possible to

minimize gold passivation.

7. Mechanism Of Cyanidation

Chemistry and electrochemistry

The most commonly used equation for the dissolution of gold in a cyanide solution,

known as Elsner‘s equation, is Gold dissolution is an electrochemical reaction in

which oxygen takes up electrons at one part of the metallic surface (the cathodic

zone), while the metal gives them up at another (the anodic zone).

According to this reaction, at low cyanide concentrations, the dissolution rate is a

function of cyanide concentration. At high cyanide concentrations, the dissolution

rate is a function of oxygen concentration. Gold dissolution is a heterogeneous

reaction that is controlled by the diffusion of both reacting species (O2 and CN_)

through the Nernst boundary layer. The rate of metal dissolution increases linearly

with increasing cyanide concentration until a maximum dissolution is reached,

beyond which there is a slight retarding effect. The dissolution rate is normally

mass-transportcontrolled in cyanide solutions with an activation energy of 8–20

kJ/mol. The formation of precipitates at the surface of gold grains is an important

aspect that determines the shape of the leaching kinetics plot.

8. Cyanide Management Plan

A key component in relation to treating cyanide is development of a sitewide

cyanide-management plan. The importance of properly developing and

administering such a plan has been highlighted by incidents at mine sites involving

the inadvertent release of cyanide to the environment. Aside from the potential for

environmental impact, such incidents broadly and negatively affect the image of the

mining industry and have led to emotional and damaging political responses, such as

the banning of cyanide in some regions.

Numerous guidance documents have been developed with regard to cyanide

management, and these documents should serve as a template for developing

site-specific cyanide management plans at mine sites.

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Implementation and adherence to a cyanide-management plan, augmented by

experienced scientific and engineering judgement, will help reduce both the number

and severity of environmental incidents involving cyanide.

The management of water and cyanide are intimately related, and development of a

cyanide-management plan should proceed in concert with development of a

water-management plan. A good cyanide-management plan will include descriptions

of how cyanide-containing solutions and slurries will be handled, stored, contained

and monitored, and in many cases the plan will also include a description of

treatment plants used to remove cyanide from solutions or slurries. At sites where

natural cyanide attenuation is important, the cyanide-management plan should

address the specifics of predicting and monitoring the effectiveness of the

attenuation processes. Decommissioning and closure are important phases in the life

cycle of cyanide management and should be addressed in the cyanide-management

plan.

9. Activated Carbon

The majority of the activated carbon used for precious metal recovery is either

granular coconut-shell carbon or peat-based extruded carbon. Important

considerations when selecting an activated carbon for use in a CIP operation include

gold-loading kinetics (activity), gold-loading capacity, elution kinetics, level to

which gold can be eluted, strength and abrasion resistance, particle-size distribution

and wet density. The properties of a particular activated-carbon sample will have a

significant impact on most aspects of the gold-recovery operation, affecting

variables such as carbon inventory, residence times, gold losses, carbon losses and

elution conditions. Therefore, due consideration must be given to the physical and

chemical properties of the virgin carbon when selecting a particular brand for

precious-metal recovery.

10. Carbon in pulp

Carbon in Pulp (CIP) is a technique for recovery of gold which has been liberated

into a cyanide solution as part of the gold cyanidation process, a gold extraction

technique.

Activated carbon acts like a sponge to aurocyanide and other complex ions in

solution. Hard carbon particles (much larger than the ore particle sizes) can be mixed

with the ore and cyanide solution mixture. The gold cyanide complex is adsorbed

onto the carbon until it comes to an equilibrium with the gold in solution. Because

http://en.wikipedia.org/wiki/Gold

http://en.wikipedia.org/wiki/Cyanide

http://en.wikipedia.org/wiki/Solution

http://en.wikipedia.org/wiki/Gold_cyanidation

http://en.wikipedia.org/wiki/Gold_extraction

http://en.wikipedia.org/wiki/Activated_carbon

http://en.wikipedia.org/w/index.php?title=Aurocyanide&action=edit&redlink=1

http://en.wikipedia.org/wiki/Solution

http://en.wikipedia.org/wiki/Carbon

http://en.wikipedia.org/wiki/Cyanide

http://en.wikipedia.org/wiki/Adsorb

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the carbon particles are much larger than the ore particles, the coarse carbon can then

be separated from the slurry by screening using a wire mesh

Flotation of gold :

The application of flotation on a reasonable scale within the gold-mining industry

commenced in the early 1930s following the introduction of watersoluble flotation

collectors (specifically xanthates and dithiophosphate collectors) that allowed

differential flotation of sulfide minerals). Prior to that time, a few gold mines in

Canada, Australia and Korea built flotation plants as the first step in the treatment of

complex and refractory gold ores. Flotation collectors on these plants were oils that

generated bulk low-grade gold concentrates, which were difficult to filter and dry.

Pre- and post-Second World War and up until late 1960s, most of the flotation

activity in the gold industry took place in Canada. During this period, Canada was

recognized as the second largest gold producer and a sizeable amount of the gold

production came from the treatment by flotation of copper–gold ores, refractory gold

ores and complex gold ores. The demand for sulfuric acid initiated by the booming

uranium industry during the late 1960s provided the catalyst for the installation of

pyrite flotation plants on numerous gold mines in South Africa. After roasting the

pyrite flotation concentrate to generate sulfur dioxide for the sulfuric acid plant, the

remaining calcine was cyanide leached to remove additional liberated gold. The

worldwide gold boom in the 1980s and 1990s created new opportunities in

Australasia and the Americas for the exploitation of medium-sized refractory gold

deposits by flotation and further treatment of the concentrates by bacterial and

pressure leaching. Many copper flotation plants around the world, and particularly

those in the Americas, have enough gold in the ore to ensure that special attention is

given to maximize the recovery of gold into the copper concentrate. A

comprehensive list of these operations, with details of the flotation reagent regimes

and circuit configurations that were in existence during the 1980s, is provided by

Bassarear. Since that time there has been a significant increase in the availability of

selective flotation collectors for gold and these are now widely in use on many large

and new copper flotation plants around the world.

4.3.1. Mineralogy

Gold occurs in a number of minerals (Harris, 1990) and the most important of these

is metallic gold and the gold metal alloys. Gold tellurides and gold–silver tellurides

are of less importance, although in particular deposits, they can account for a

significant proportion of the gold content. Aurostibite [AuSb2] and maldonite

[Au2Bi] are rare minerals but are present in some gold deposits. Aurocupride

[AuCu] is found in some primary copper ores. Gold particles in an ore deposit will

http://en.wikipedia.org/wiki/Slurry

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vary in size from large nuggets to particles locked in the crystal lattice of sulfide

minerals. These sulfide minerals are referred to as gold carrier minerals and contain

trace to minor amounts of gold Often it is found that gold ores are refractory due to

the small size of the gold particles in the sulfides and concentration by flotation is

required, followed either by roasting, bacterial leaching or pressure leaching to

liberate the gold prior to cyanidation.

General aspects of gold flotation

Most of the reported fundamental work on the flotation of gold has been conducted

using high-purity gold and gold–silver alloys with the purpose of determining

collector–gold interactions and the nature of adsorption of collector ions or

molecules onto the gold surface. In addition, some work has been conducted to

decide whether or not pure gold has a natural hydrophobicity and hence some degree

of natural floatability. Flotation research work has been conducted on naturally

occurring native gold particles recovered from placer deposits and on gold particles

selected from lode deposits. The flotation characteristics of gold or gold minerals

found in refractory sulfide and copper ores have not been described in detail in the

literature. The sparse distribution of discrete gold minerals and particles, as well as

their exceedingly low concentration in ores, are the principal reasons for the lack of

fundamental work on gold flotation. A great deal of work has been reported on

specific ores, but such studies rarely distinguish between the flotation of native gold

and other gold minerals. Flotation of gold ores covers a broad field and it is a rather

difficult subject to generalize on. Most problems in gold ore flotation are not

connected with floating metallic gold. The flotation recovery of free gold

(throughout the text free gold is synonymous with liberated gold) is largely affected

by physical constraints such as the shape and size of the gold particles and the

stability of the froth. It is a generally accepted fact that liberated gold finer than about

150 mm floats readily with most collectors and in particular xanthates and

dithiophosphates. When free gold is floated with other sulfide minerals the extent of

bubble loading of sulfide particles may provide a barrier towards the attachment of

free gold, thereby reducing flotation performance. Research investigations have until

recently typically focused on the individual flotation behaviour of each gold-bearing

mineral in synthetic mixtures and not on mixtures of sulfide minerals in ‗real‘ ores.

In the flotation process, the main chemical effects are reagent type and pulp pH.

Recently, there has been a need to operate circuits at moderate pH levels, to improve

separation efficiencies when treating complex low-grade ores, to reduce costs of

reagents, to develop reagents that are stable over a wide pH range and to take

advantage of the synergistic benefit of mixture collector systems. This has led to

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ever-increasing research to develop new collectors and mixtures for the flotation of

gold-bearing.

Surface characteristics of pure gold

Several research papers have dealt with the hydrophilic and hydrophobic nature of

pure metallic gold and whether either a zero or a finite contact angle is observed.

Some workers measured contact angles on gold; while others did not. The

discrepancy between the findings of zero and of high contact angles on a pure gold

surface appeared to be due to both the presence of residual polishing agents and

organic contaminants. It is now generally accepted that the surface of pure clean gold

is hydrophilic and displays a zero contact angle. The hydrophilicity of gold is due to

the high Hamaker constant and is a result of the strong dispersion attraction forces

for water.

Collector less Flotation of Naturally Occurring

GOLD

Numerous references are to be found in the literature on the observed skin flotation

of gold, especially during the recovery of gold by gravity separation .On this

evidence gold was presumed to be naturally hydrophobic. As discussed above,

naturally occurring native gold surfaces are usually found to be hydrophobic; this is a

result of the contamination of the gold surface by organic compounds. Untarnished

gold of the appropriate particle size has been found to readily float with only a . The

earliest recorded laboratory work on gold flotation found that gold floated in the

presence of frother only, but not if its particle size was too large or if reagents such as

calcium oxide or sodium sulfide were added to the pulp. Gold can also be rendered

hydrophobic by the deposition of sulfur on the surface.

4.3.5.Collectors in Gold Flotation

Collector flotation of naturally occurring, placer and liberated gold

Gold hydrophobicity is enhanced by the addition of flotation collectors and no

flotation plant relies solely on the natural floatability of gold for its recovery.

Naturally occurring or free (liberated) gold is optimally recovered in a flotation

circuit at natural or near-natural pulp pH values and with the addition of small

amounts of collector. Inherently, naturally floating minerals float fast kinetically

Flotation tests on placer gold .showed that fine placer gold typically floated readily

with common sulfhydryl collectors and common frothers at natural pH without the

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addition of any special regulating reagents. Gold flotation recoveries ranged from 78

to 99%.

Flotation collectors for gold and gold carriers

Flotation with xanthate collectors involves the anodic oxidation of the collector that

may involve sub-processes such as metal xanthate formation, chemisorption of the

xanthate ion and oxidation of the xanthate to form dixanthogen. These adsorb onto

mineral surfaces, rendering the mineral hydrophobic. It is generally accepted that the

xanthate species responsible for the flotation of free gold is dixanthogen. This is a

neutral oil that will adsorb onto the surface of any naturally hydrophobic solid,

rendering it floatable. Dixanthogen may form on gold by either the application of an

applied potential or by a mixed potential mechanism in a pulp that involves the

reduction of oxygen. Studies have shown that the development of a finite contact

angle and the onset of flotation of gold particles occur at a potential close to that of

dixanthogen formation. The longer-chain xanthates are more readily oxidized,

generating dixanthogen at lower potentials. An increase in thiol chain length

increases the maximum contact angle, thereby increasing the hydrophobicity of the

surface species. Both these attributes favour the use of longer chain xanthates, such

as potassium amyl xanthate (PAX) for the flotation of free gold.

It is quite common to encounter silver and other precious metals forming alloys with

native gold. The positive effect that silver has on gold floatability was first

recognized in experiments using plates of pure gold, silver and gold–silver alloys.

The adsorption of ethyl xanthate on silver is generally thought to take place through

an electrochemical mechanism of metal xanthate formation on the surfaces . For

ethyl xanthate, the presence of silver in gold leads to silver xanthate formation at a

potential proportionately lower than for dixanthogen formation on pure gold . As a

consequence, the flotation of gold–silver alloys can be achieved at potentials

considerably lower than that for gold. Xanthate ions chemisorb on silver at potentials

below the region at which silver xanthate deposits. Chemisorbed ethyl xanthate

results in finite contact angles on silver surfaces and the initiation of flotation

appears to result from the chemisorptions process. For more rapid flotation

dixanthogen may play a supporting role. The chemisorbed sub-monolayer could be

important in retaining the dixanthogen at the gold surface through hydrophobic

interactions between the adsorbate and the bulk phase. The xanthogen formates are

produced by reacting alkyl chloroformate with xanthate salts. They are stable in

acidic conditions unlike the xanthates from which they are formed and are stable in

the pH range of 5–10.5. The formates appear to have superior pyrite rejection

properties compared to xanthates and dithiophoshate .

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Frothers in Gold Flotation

The strength and stability of the froth is important when floating free gold. There

appears to be a preference for polyglycol ether-based frothers on most gold plants in

combination with one or other frothers. When selectivity is required or, in the case of

copper–gold ores, where a copper concentrate is sold to a smelter, a weaker frother

such as methyl isobutyl carbinol (MIBC) is preferred. The choice of a particle

size-balanced frother is also an important consideration in gold flotation as this

promotes composite particle recovery in the scavenger flotation circuit. As a rule, the

glycol or polypropylene glycol methyl ether frothers are ideal for this application.

The blended interfroth frothers have found wide acceptance on Australian gold

plants and the base reagent is an alkyl aryl ester.

4.3.8.Activators in Gold Flotation

Activation implies improved floatability of a mineral after the addition of a soluble

base metal salt or sulfidizer. It is generally thought that the metal or sulfide ion

adsorbs onto the mineral surface thus changing its surface chemical properties. In

this way, the flotation response can be improved and/or the pH range of flotation for

the mineral can be extended, the rates of flotation increased and selectivity

improved.

Sulfidization

The application of sulfidizers (sodium sulfide and sodium hydrosulfide) to enhance

the flotation of oxidized ores is well. The first detailed laboratory study of the

influence of sodium sulfide on the flotation of gold-bearing ores was undertaken in

the mid-1930s. The outcome from this study was that, in general, sodium sulfide

retards the flotation of gold, although for some ores there was benefit in its addition.

Similar comments are to be found in the literature since that time. Sulfide ions appear

to act as flotation activators at low concentrations (less than 10_5M) and as a strong

depressant at concentrations above 10_5 M. The addition of sulfide ions converts

some coatings on mineral surfaces in sulfides and subsequent xanthate addition will

promote flotation. For successful activation, the sulfide activator should be added

slowly and at starvation quantities.

A recent study of a number of flotation plants found that floated gold grains in the

concentrate had a greater concentration of silver and sulfur than the gold remaining

in the tailings, implying that they assisted gold flotation. Laboratory flotation tests on

a flotation-plant tailing sample using NaHS and Silver ions recovered 30–45% of the

unfloated gold and NaSH provided the best results. Application of this information at

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plant scale resulted in a 7% increase in gold recovery at the Los Pelambres Mine in

Chile. Rejected gold particles at this mine were found to be coated with lead

carbonate.

A recent innovation, the Controlled Potential Sulfidization (CPS) process, where the

pulp potential is controlled has been successfully applied at a number of copper–gold

ore flotation plants in Australia In this process, sodium sulfide or sodium hydrogen

sulfide is added and the solution potential is controlled at about _450mV prior to

flotation.

4.3.9. Depression of Gold In Flotation

Depressants for native gold that are usually introduced during the flotation process

include compounds such as calcium ions, chloride ions, calcium carbonate, cyanide,

sodium silicate, sodium sulfite, ferric and heavy metal ions, tannin and related

compounds, starch and other organic depressants and many others . All of these may

competitively adsorb on the gold surface thus preventing the adsorption of the

collector(s) added. It has also been suggested that the ferric ions, which would be in

the form of hydrated oxides, may act as a physical barrier between the air bubble and

gold surface but this effect is reversed simply by washing with water However,

flotation of native gold often proceeds satisfactorily in the presence of many of these

compounds. In general, the results reported by different authors are not in good

agreement (Allan and Woodcock, 2001). It is likely that other components in

solution or on the surface of the gold that were not measured provide the answer for

the different outcomes. Lime cannot be considered as just a pH modifier and studies

have shown that calcium is strongly adsorbed on sulfide minerals and gold at pH

values at and above 10. This adsorption is enhanced if excess sulfate in the pulp

promotes calcium-sulfate coatings on particles.

Desorption of calcium from the surface by reducing the pH can be assisted by the use

of specific calcium-complexing ions such as polyphosphate. Furthermore, if the

calcium release is attempted while adding excess activator, then a hydrophilic

hydroxide coating can result. Metals ions introduced from the circuit water, or from

soluble metal ions in the ore, may adsorb and nucleate as hydroxide coatings on all

particle surfaces, thus inhibiting collector adsorption. The recommended method of

flotation treatment is to operate at as low a pH value as practical, avoid rapid

increases in pH, add activator slowly or condition separately and keep the tailings

dam at a pH of minimum solubility.

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4.3.10 Flotation of Gold and Gold-Bearing Minerals

Differential flotation of natural and liberated gold

Experience has shown that free gold particles can be recovered selectively against

pyrite, by keeping the gold particle surfaces as clean as possible of organic species

and by removing any adhering slime particles This can be achieved with the use of

little or no pH regulators, only small dosages of collectors and suitable frother to

stabilize the froth, and possibly a small amount of dispersant. Selectivity for gold

against pyrite was found to be enhanced in the presence of collectors such as alkoxy

or phenoxy carbonyl alkyl thionocarbamates, dialkyl or diaryl monothiophosphates

and monothiophosphinates, glyoxalidine and aminothiophenols.

Monothiophosphorous acids have been shown to be able to float gold

Selectively from base metal sulfides.

The use of hydrogen peroxide as an oxidizing agent in the selective flotation of gold

from pyrite with PAX has been demonstrated at laboratory scale. Hydrogen peroxide

addition by itself rendered both the gold and pyrite surfaces hydrophobic. The

addition of xanthate converted the gold surface into a fairly hydrophobic condition,

whereas the pyrite was still hydrophilic at pH values of 10 and higher. On most

flotation plants, there is a tendency to treat sulfide ores containing free gold as

though the gold is associated in a massive or complex sulfide mineral matrix. This

leads to high dosage levels of collector and activator addition. In this application,

xanthate adsorption on both sulfides and gold makes selective flotation rather

difficult due to the formation of dixanthogen on both the gold and sulfide surfaces.

4.3.12. Flotation of gold-carrying iron sulfides

Pyrite is known to be oleophilic when the surface is free of oxidation products but it

is nevertheless necessary to use a collector to float pyrite. Thiol collectors are the

most commonly used Pyrite can be recovered optimally in either acidic or alkaline

conditions. Xanthates are the primary collectors on the alkaline side and MBT is

used when floating in an acidic medium. Blends of different collectors are

particularly effective for recovering pyrite and mixtures of xanthate, thiocarbamates

and xanthate and xanthate mixed with MBTs have been successfully used.

Dithiophosphates are usually used as secondary collectors and on their own are

reported to be selective against pyrite and to a lesser extent against arsenopyrite.

Aminebased collectors are capable of floating cyanide-leached pyrite without the use

of acid pre-conditioning.

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Arsenopyrite has very similar properties to pyrite and the flotation conditions for its

recovery are similar to pyrite. Arsenopyrite is susceptible to the formation of

oxidation products on its surface. Under oxidizing conditions, ahigh concentration of

PAX is required in order to achieve high arsenopyrite recovery. The reason for this is

not clear but the low rate of arsenopyrite floatability could be attributed to the

formation of iron-oxide species on the surface of the arsenopyrite. Oxidation

products such as ferric and ferrous ions are also present in the pulp and depending on

the chemical conditions these promote the formation of dixanthogen and bulk

ferric-xanthate compounds in solution. Pyrrhotite floats readily in acid and neutral

pH ranges. Surface coatings in the alkaline regions may result in depressed flotation

recovery, while collector regimes are similar to those for pyrite flotation. Pyrrhotite

oxidizes readily and this makes it more difficult to float than arsenopyrite. Oxidation

products in solution create flotation problems similar to those mentioned above for

arsenopyrite. The floatability of marcasite appears to be variable as it has been

shown to float more readily than pyrite, while more recently.

4.3.14. Flotation of copper– gold ores

For the bulk flotation of copper minerals and gold from supergene copper ores, it is

normal practice to add a xanthate as the primary collector and dithiophosphate as a

secondary collector. This combination gives satisfactory results in respect of copper

concentrate grade and copper and gold recoveries. When pyrite is present in the ore,

the choice of collectors will generally depend on the ratio of copper to pyrite in the

feed. Dithiophosphate collectors are found to be more selective against pyrite and

better gold collectors than xanthate . With low pyrite contents in the feed, the

ultimate gold recovery is dependent on the type of xanthate selected, longer

carbon-chain lengths achieving higher gold recovery. Ores containing large amounts

of pyrite require collectors that are selective towards both the copper minerals and

free gold. For these ores, the preferred collectors are the dithiophosphates and the

new generation of ‗gold‘ collectors. Recent laboratory testwork on a copper–gold ore

containing pyrite showed that Aerofloat 7249 achieved the highest gold . Aerofloat

6697 provided the best gold concentrate grade at pH values less than 11.5 while

Aerofloat 7249 and Aerofloat 208 were better at pH values above 11.5 and gold

grades in excess of 250 g/t were generated. At industrial scale, the use of Aerofloat

7249 provided a 4.5% improved gold recovery at the Freeport Copper Mine, while

the addition of 3477 in the cleaner scavenger circuit improved the recovery of

tarnished flaky gold at the Kemess Copper–Gold Mine in Canada.

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4.3.15. Particle size and shape in flotation

It is well known that particle size is an important parameter in flotation and that size

limits exist at which minerals will and will not float. The high particle-density of

gold and its malleable and ductile properties that favour the propagation of platy

particles, further compound this effect. Platy/flaky particles are formed in the

treatment process, particularly in grinding, or during transportation events in nature .

During these events, some gold particles are impregnated with nonfloatable particles

(Taggart, 1945; Pevzner et al., 1966), inhibiting flotation. Passivation of a

gold-particle surface may also occur after considerable hammering by steel

grinding-media . On the other hand, it is postulated that the surface of the gold could

become more active and therefore more floatable due to work hardening. It has been

suggested that the practical particle size limits for gold flotation are around 5–200

mm. Particles as small as 3 mm have been floated at laboratory scale), while actual

measurements indicate that the flotation performance on many gold plants decrease

rapidly below 10 mm (Chryssoulis, 2004). At the coarse end, gold particles as large

as 300 mm (Leaver and Woolf, 1934b) and 700 mm have been floated in laboratory

flotation cells under specific operating conditions and high collector additions (Lins

and Adamian, 1993). Flotation of 590 mm gold particles has been reported on an

industrial scale with ‗unit‘ flotation cells (Leaver and Woolf, 1934). Pulp density and

aeration rates influence flotation-cell pulp hydrodynamics and are important

parameters in extending the particle-size limits of gold flotation. There is conflicting

commentary on the best pulp density for gold particle flotation, both a high

pulp-density (Leaver and Woolf, 1934a) and a low pulp-density being

recommended.

4.3.16. FLOTATION CIRCUITS

Flotation circuit configuration on most gold mines can be divided into a number of

categories, viz. open circuits with no cleaning at all, and open and closed circuits

with single stage and two stages of cleaning. Open circuits have the advantage of no

feedback from the effects of non-steady-state operation and therefore are inherently

more stable than the closed-circuit configuration.

Closed and open-circuit flotation cleaning is used on gold mines where high-grade

concentrates are required for roasting and smelting. Under these conditions, it is

difficult to maintain very high gold and sulfide flotation recoveries, while also

producing an acceptable grade of concentrate. Where there is no constraint on

concentrate quality, high gold and sulfide flotation recoveries are achievable to the

extent that a discardable gold flotation tail is possible. Cleaning-circuit

configuration, either single or two stages of cleaning, and cleaner residence time are

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related to the particle size of the sulfides in the flotation feed and also the presence or

absence of floatable gangue components. Unit flotation cells and the more recent

Flash flotation cells are installed in grinding circuits with the purpose of improving

the overall flotation recovery of free gold . The aim is to remove as much of the free

gold contained in the circulating load of the grinding mill before it is overground or

is covered with coatings of iron, sulfide or other coatings that will lower flotation

recoveries. Improved overall gold flotation recoveries of 2–10% have been quoted

(Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990).

Furthermore, the inclusion of Unit and Flash flotation cells will generally provide

better flotation stability and performance. Improved overall gold flotation recoveries

from 3% to 10%

Slime coatings and floatable non-sulfide gangue

The deleterious impact of clay slimes on gold flotation is well. The failure of free

gold and sulfide minerals to float has at times been shown to be related to the

presence of coatings of colloidal or near-colloidal gangue or silicate material

adhering to the mineral surface.

These coatings are formed under pulp conditions in which the sulfide particles and

silicate particles are oppositely charged. Gangue minerals that are known to cause

problems include talcose and carbonaceous minerals, bentonite clay, goethite

[FeO(OH)] iron oxide and manganese slimes, pyrophyllite [AlSi2O5OH] and

carbonates . Slime coatings are controlled by the use of gangue-dispersing agents.

Sodium silicate is widely used for this purpose and is most effective when the

alkalinity is carefully controlled. Sodium sulfide has also been found to be an

effective dispersing agent. In addition to coating the mineral surface, the gangue

particles may coat the bubble surface, affecting the ability of any gold and sulfide

particles to attach to the air bubbles. Other more recent remedies to overcome the

problem of slime coatings have included physical methods such as removal of the

slimes by.

Organic compounds of high molecular weight that maintain a state of dispersion of

deleterious slime components by forming wettable coatings on the gangue particles

are used for much the same purpose as the inorganic dispersing agents. These

organic compounds are referred to as organic gangue depressants. Typical examples

are glue, starch, dextrin, gum arabic, carboxymethylcellulose and the more recent

modified-guar gums. Selection of the correct depressant type and dosage is critical,

as an overdose results in both loss of free gold and sulfides that contain gold. The

anionic polymers generally have a negligible depressant capability on sulfide

minerals while the cationic polymers are capable of acting as sulfidemineral

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depressants .The combination of a collector and depressant is also important since in

the flotation of pyrite, for example, guar gum will have a more adverse effect when

used with MBT than with xanthate . An alternative approach that has had some

success is to add small quantities of frother to the pulp and selectively float the

talcaceous minerals prior to removing the bulk sulfide concentrate. These talc

concentrates may contain up to 30%–40% of the gold contained in the feed and this

concentrate can either be cyanide leached separately or recombined with the sulfide

tailing prior to cyanidation .

Porphyry copper–gold ores usually contain some gangue components that are highly

floatable and contaminate the copper concentrate. Maintenance of a high

copper-concentrate grade requires that gangue depressants be used. Silicates, guards

and carboxymethylcellulose are the common depressants applied in the copper

industry.

Carbonaceous and graphitic minerals are soft and flaky, and easily broken down

during grinding. During flotation, the carbon floats readily owing to its fine grain

size, natural hydrophobicity, platy nature and low density. Graphitic carbon and

clays can be the cause of poor gold recovery on many refractory gold or flotation

plants.

Natural metal and organic coatings on gold

Most coatings on mineral surfaces are detrimental to flotation, but in some cases, the

effects can be overcome. Many types of surface coatings have been reported to

occur on native gold particles. Perhaps the most difficult coatings to cope with are

hydrated iron oxides. The surfaces of gold particles can become coated naturally

with precipitates of iron, from oxidized sulfides in an orebody or from rusting iron,

such as iron grinding media, as first reported by Head (1936). Gold from placer

deposits heavily stained or coated by iron oxides or impregnated by hydrophilic

minerals is not easily floated. Tarnished gold has been found also to have a markedly

higher mercury content compared to ‗shiny‘ gold .Gold particles coated with

manganese dioxide have also been reported . Some gold flotation pulps may contain

humic and tannin substances) and sulfide ions from sulfide mineral that are reported

to impact on gold flotation. Humic acid has been found to be only marginally

deleterious to gold flotation and some naturally occurring organic coatings can be

removed by conditioning with sodium hydroxide or acid solution .

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Factors For Consideration In Refractory Process Selection

Based on GRD Minproc‘s experience on numerous projects, the following factors

are considered to be of importance in selecting a process for treatment of refractory

gold ores. All factors should be taken into account at an early stage and process

options should be kept open for as long as possible due to the potential for

unforeseen issues to impact on the economics of a particular process.

Refractory gold ores

Successful concentration of gold in refractory sulfide ores is almost exclusively

dependent on the association of the gold with the sulfides . Refractory gold ores

commonly contain free gold, sub-microscopic gold, carbonaceous material, base

metals, pyrite, marcasite, arsenopyrite and pyrrhotite . Clays and graphitic carbon are

the most troublesome accessory components in some of these ores, as far as gold

concentration is concerned. Arsenopyrite has very similar properties to pyrite and the

flotation conditions for its recovery are similar to pyrite. Arsenopyrite is marginally

less hard and more brittle than pyrite and pyrrhotite . During milling, the

arsenopyrite is therefore subject to more overgrinding than pyrite and pyrrhotite. The

difference in the recovery of these minerals is due not only to the difference in the

surface chemical properties of the particle but also to the difference in their overall

size distribution

.

Reference

Advances in gold ore processing

A method for leaching or dissolving gold from ore

environmental management in mining

Gold Project 2011

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Transcript of Gold Project 2011